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  • How to create a SOAP REQUEST using ASP.NET (VB) without using Visual

    - by user311691
    Hi all , I urgently need your help . I am new to consuming a web service using SOAP protocol. I have been given a demo webservice URL which ends in .WSDL and NOT .asml?WSDL. The problem is I cannot add a web reference using Visual studio OR Disco.exe or Wsdl.exe - This webservice has been created on a java platform and for security reasons the only way to make a invoke the webservice is at runtime using SOAP protocol IN asp.net (VB). I I have created some code but cannot seem to send the soap object to the receiving web service. If I could get a solution with step by step instructions on how I can send a SOAP REQUEST. Below is my code and all am trying to do is send a SOAP REQUEST and receive a SOAP RESPONSE which I will display in my browser. <%@ page language="vb" %> <%@ Import Namespace="System.Data"%> <%@ Import Namespace="System.Xml"%> <%@ Import Namespace="System.Net"%> <%@ Import Namespace="System.IO"%> <%@ Import Namespace="System.Text"%> <script runat=server> Private Sub Page_Load() Dim objHTTPReq As HttpWebRequest Dim WebserviceUrl As String = "http://xx.xx.xx:8084/asy/wsdl/asy.wsdl" objHTTPReq = CType(WebRequest.Create(WebserviceUrl), HttpWebRequest) Dim soapXML As String soapXML = "<?xml version='1.0' encoding='utf-8'?>" & _ " <soap:Envelope xmlns:xsi='http://www.w3.org/2001/XMLSchema-instance'" & _ " xmlns:xsd='http://www.w3.org/2001/XMLSchema'"& _ " xmlns:soap='http://schemas.xmlsoap.org/soap/envelope/' >"& _ " <soap:Body> "& _ " <validatePaymentData xmlns='http://asybanks.webservices.asycuda.org'> " & _ " <bankCode>"& bankCode &"</bankCode> " & _ " <PaymentDataType>" & _ " <paymentType>"& payment_type &"</paymentType> " & _ " <amount>"& ass_amount &"</amount> " & _ " <ReferenceType>" & _ " <year>"& year &"</year> " & _ " <customsOfficeCode>"& station &"</customsOfficeCode> " & _ " </ReferenceType>" & _ " <accountNumber>"& zra_account &"</accountNumber> " & _ " </PaymentDataType> " & _ " </validatePaymentData> " & _ " </soap:Body> " & _ " </soap:Envelope> " objHTTPReq.Headers.Add("SOAPAction", "http://asybanks.webservices.asycuda.org") objHTTPReq.ContentType = "text/xml; charset=utf-8" objHTTPReq.ContentLength = soapXML.Length objHTTPReq.Accept = "text/xml" objHTTPReq.Method = "POST" Dim objHTTPRes As HttpWebResponse = CType(objHTTPReq.GetResponse(), HttpWebResponse) Dim dataStream As Stream = objHTTPRes.GetResponseStream() Dim reader As StreamReader = new StreamReader(dataStream) Dim responseFromServer As String = reader.ReadToEnd() OurXml.text = responseFromServer End Sub </script> <html xmlns="http://www.w3.org/1999/xhtml"> <head runat="server"> <title> XML TRANSACTION SIMULATION - N@W@ TJ </title> </head> <body> <form id="form1" runat="server"> <div> <p>ZRA test Feedback:</p> <asp:label id="OurXml" runat="server"/> </div> </form> </body> </html> the demo webservice looks like this: <?xml version="1.0" encoding="UTF-8" ?> - <!-- WEB SERVICE JAVA DEMO --> - <definitions targetNamespace="http://asybanks.webservices.asycuda.org" xmlns="http://schemas.xmlsoap.org/wsdl/" xmlns:apachesoap="http://xml.apache.org/xml-soap" xmlns:soap="http://schemas.xmlsoap.org/wsdl/soap/" xmlns:xs="http://www.w3.org/2001/XMLSchema" xmlns:y="http://asybanks.webservices.asycuda.org"> - <types> - <xs:schema elementFormDefault="qualified" targetNamespace="http://asybanks.webservices.asycuda.org" xmlns="http://www.w3.org/2001/XMLSchema"> SOME OTHER INFORMATION AT THE BOTTOM <soap:address location="http://xx.xx.xx:8084/asy/services/asy" /> </port> </service> </definitions> From the above excerpt of the wsdl url webservice, I am not sure which namespace to use for soapACTION - please advise.... Please if you could comment every stage of a soap request and provide a working demo - I would be most grateful as I would be learning rather than just assuming stuff :)

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  • Trouble with copying dictionaries and using deepcopy on an SQLAlchemy ORM object

    - by Az
    Hi there, I'm doing a Simulated Annealing algorithm to optimise a given allocation of students and projects. This is language-agnostic pseudocode from Wikipedia: s ? s0; e ? E(s) // Initial state, energy. sbest ? s; ebest ? e // Initial "best" solution k ? 0 // Energy evaluation count. while k < kmax and e > emax // While time left & not good enough: snew ? neighbour(s) // Pick some neighbour. enew ? E(snew) // Compute its energy. if enew < ebest then // Is this a new best? sbest ? snew; ebest ? enew // Save 'new neighbour' to 'best found'. if P(e, enew, temp(k/kmax)) > random() then // Should we move to it? s ? snew; e ? enew // Yes, change state. k ? k + 1 // One more evaluation done return sbest // Return the best solution found. The following is an adaptation of the technique. My supervisor said the idea is fine in theory. First I pick up some allocation (i.e. an entire dictionary of students and their allocated projects, including the ranks for the projects) from entire set of randomised allocations, copy it and pass it to my function. Let's call this allocation aOld (it is a dictionary). aOld has a weight related to it called wOld. The weighting is described below. The function does the following: Let this allocation, aOld be the best_node From all the students, pick a random number of students and stick in a list Strip (DEALLOCATE) them of their projects ++ reflect the changes for projects (allocated parameter is now False) and lecturers (free up slots if one or more of their projects are no longer allocated) Randomise that list Try assigning (REALLOCATE) everyone in that list projects again Calculate the weight (add up ranks, rank 1 = 1, rank 2 = 2... and no project rank = 101) For this new allocation aNew, if the weight wNew is smaller than the allocation weight wOld I picked up at the beginning, then this is the best_node (as defined by the Simulated Annealing algorithm above). Apply the algorithm to aNew and continue. If wOld < wNew, then apply the algorithm to aOld again and continue. The allocations/data-points are expressed as "nodes" such that a node = (weight, allocation_dict, projects_dict, lecturers_dict) Right now, I can only perform this algorithm once, but I'll need to try for a number N (denoted by kmax in the Wikipedia snippet) and make sure I always have with me, the previous node and the best_node. So that I don't modify my original dictionaries (which I might want to reset to), I've done a shallow copy of the dictionaries. From what I've read in the docs, it seems that it only copies the references and since my dictionaries contain objects, changing the copied dictionary ends up changing the objects anyway. So I tried to use copy.deepcopy().These dictionaries refer to objects that have been mapped with SQLA. Questions: I've been given some solutions to the problems faced but due to my über green-ness with using Python, they all sound rather cryptic to me. Deepcopy isn't playing nicely with SQLA. I've been told thatdeepcopy on ORM objects probably has issues that prevent it from working as you'd expect. Apparently I'd be better off "building copy constructors, i.e. def copy(self): return FooBar(....)." Can someone please explain what that means? I checked and found out that deepcopy has issues because SQLAlchemy places extra information on your objects, i.e. an _sa_instance_state attribute, that I wouldn't want in the copy but is necessary for the object to have. I've been told: "There are ways to manually blow away the old _sa_instance_state and put a new one on the object, but the most straightforward is to make a new object with __init__() and set up the attributes that are significant, instead of doing a full deep copy." What exactly does that mean? Do I create a new, unmapped class similar to the old, mapped one? An alternate solution is that I'd have to "implement __deepcopy__() on your objects and ensure that a new _sa_instance_state is set up, there are functions in sqlalchemy.orm.attributes which can help with that." Once again this is beyond me so could someone kindly explain what it means? A more general question: given the above information are there any suggestions on how I can maintain the information/state for the best_node (which must always persist through my while loop) and the previous_node, if my actual objects (referenced by the dictionaries, therefore the nodes) are changing due to the deallocation/reallocation taking place? That is, without using copy?

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  • creating a color coded time chart using colorbar and colormaps in python

    - by Rusty
    I'm trying to make a time tracking chart based on a daily time tracking file that I used. I wrote code that crawls through my files and generates a few lists. endTimes is a list of times that a particular activity ends in minutes going from 0 at midnight the first day of the month to however many minutes are in a month. labels is a list of labels for the times listed in endTimes. It is one shorter than endtimes since the trackers don't have any data about before 0 minute. Most labels are repeats. categories contains every unique value of labels in order of how well I regard that time. I want to create a colorbar or a stack of colorbars (1 for eachday) that will depict how I spend my time for a month and put a color associated with each label. Each value in categories will have a color associated. More blue for more good. More red for more bad. It is already in order for the jet colormap to be right, but I need to get desecrate color values evenly spaced out for each value in categories. Then I figure the next step would be to convert that to a listed colormap to use for the colorbar based on how the labels associated with the categories. I think this is the right way to do it, but I am not sure. I am not sure how to associate the labels with color values. Here is the last part of my code so far. I found one function to make a discrete colormaps. It does, but it isn't what I am looking for and I am not sure what is happening. Thanks for the help! # now I need to develop the graph import numpy as np from matplotlib import pyplot,mpl import matplotlib from scipy import interpolate from scipy import * def contains(thelist,name): # checks if the current list of categories contains the one just read for val in thelist: if val == name: return True return False def getCategories(lastFile): ''' must determine the colors to use I would like to make a gradient so that the better the task, the closer to blue bad labels will recieve colors closer to blue read the last file given for the information on how I feel the order should be then just keep them in the order of how good they are in the tracker use a color range and develop discrete values for each category by evenly spacing them out any time not found should assume to be sleep sleep should be white ''' tracker = open(lastFile+'.txt') # open the last file # find all the categories categories = [] for line in tracker: pos = line.find(':') # does it have a : or a ? if pos==-1: pos=line.find('?') if pos != -1: # ignore if no : or ? name = line[0:pos].strip() # split at the : or ? if contains(categories,name)==False: # if the category is new categories.append(name) # make a new one return categories # find good values in order of last day newlabels=[] for val in getCategories(lastDay): if contains(labels,val): newlabels.append(val) categories=newlabels # convert discrete colormap to listed colormap python for ii,val in enumerate(labels): if contains(categories,val)==False: labels[ii]='sleep' # create a figure fig = pyplot.figure() axes = [] for x in range(endTimes[-1]%(24*60)): ax = fig.add_axes([0.05, 0.65, 0.9, 0.15]) axes.append(ax) # figure out the colors to use # stole this function to make a discrete colormap # http://www.scipy.org/Cookbook/Matplotlib/ColormapTransformations def cmap_discretize(cmap, N): """Return a discrete colormap from the continuous colormap cmap. cmap: colormap instance, eg. cm.jet. N: Number of colors. Example x = resize(arange(100), (5,100)) djet = cmap_discretize(cm.jet, 5) imshow(x, cmap=djet) """ cdict = cmap._segmentdata.copy() # N colors colors_i = np.linspace(0,1.,N) # N+1 indices indices = np.linspace(0,1.,N+1) for key in ('red','green','blue'): # Find the N colors D = np.array(cdict[key]) I = interpolate.interp1d(D[:,0], D[:,1]) colors = I(colors_i) # Place these colors at the correct indices. A = zeros((N+1,3), float) A[:,0] = indices A[1:,1] = colors A[:-1,2] = colors # Create a tuple for the dictionary. L = [] for l in A: L.append(tuple(l)) cdict[key] = tuple(L) # Return colormap object. return matplotlib.colors.LinearSegmentedColormap('colormap',cdict,1024) # jet colormap goes from blue to red (good to bad) cmap = cmap_discretize(mpl.cm.jet, len(categories)) cmap.set_over('0.25') cmap.set_under('0.75') #norm = mpl.colors.Normalize(endTimes,cmap.N) print endTimes print labels # make a color list by matching labels to a picture #norm = mpl.colors.ListedColormap(colorList) cb1 = mpl.colorbar.ColorbarBase(axes[0],cmap=cmap ,orientation='horizontal' ,boundaries=endTimes ,ticks=endTimes ,spacing='proportional') pyplot.show()

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  • C -Segmentation fault !

    - by FILIaS
    It seems at least weird to me... The program runs normally.But after I call the enter() function for the 4th time,there is a segmentation fault!I would appreciate any help. With the following function enter() I wanna add user commands' datas to a list. [Some part of the code is already posted on another question of me, but I think I should post it again...as it's a different problem I'm facing now.] /* struct for all the datas that user enters on file*/ typedef struct catalog { char short_name[50]; char surname[50]; signed int amount; char description[1000]; struct catalog *next; }catalog,*catalogPointer; catalogPointer current; catalogPointer head = NULL; void enter(void) //user command: i <name> <surname> <amount> <description> { int n,j=2,k=0; char temp[1500]; char *short_name,*surname,*description; signed int amount; char* params = strchr(command,' ') + 1; //strchr returns a pointer to the 1st space on the command.U want a pointer to the char right after that space. strcpy(temp, params); //params is saved as temp. char *curToken = strtok(temp," "); //strtok cuts 'temp' into strings between the spaces and saves them to 'curToken' printf("temp is:%s \n",temp); printf("\nWhat you entered for saving:\n"); for (n = 0; curToken; ++n) //until curToken ends: { if (curToken) { short_name = malloc(strlen(curToken) + 1); strncpy(short_name, curToken, sizeof (short_name)); } printf("Short Name: %s \n",short_name); curToken = strtok(NULL," "); if (curToken) { surname = malloc(strlen(curToken) + 1); strncpy(surname, curToken,sizeof (surname)); } printf("SurName: %s \n",surname); curToken = strtok(NULL," "); if (curToken) { //int * amount= malloc(sizeof (signed int *)); char *chk; amount = (int) strtol(curToken, &chk, 10); if (!isspace(*chk) && *chk != 0) fprintf(stderr,"Warning: expected integer value for amount, received %s instead\n",curToken); } printf("Amount: %d \n",amount); curToken = strtok(NULL,"\0"); if (curToken) { description = malloc(strlen(curToken) + 1); strncpy(description, curToken, sizeof (description)); } printf("Description: %s \n",description); break; } if (findEntryExists(head, surname,short_name) != NULL) //call function in order to see if entry exists already on the catalog printf("\nAn entry for <%s %s> is already in the catalog!\nNew entry not entered.\n",short_name,surname); else { printf("\nTry to entry <%s %s %d %s> in the catalog list!\n",short_name,surname,amount,description); newEntry(&head,short_name,surname,amount,description); printf("\n**Entry done!**\n"); } // Maintain the list in alphabetical order by surname. } catalogPointer findEntryExists (catalogPointer head, char num[],char first[]) { catalogPointer p = head; while (p != NULL && strcmp(p->surname, num) != 0 && strcmp(p->short_name,first) != 0) { p = p->next; } return p; } catalogPointer newEntry (catalog** headRef,char short_name[], char surname[], signed int amount, char description[]) { catalogPointer newNode = (catalogPointer)malloc(sizeof(catalog)); catalogPointer first; catalogPointer second; catalogPointer tmp; first=head; second=NULL; strcpy(newNode->short_name, short_name); strcpy(newNode->surname, surname); newNode->amount=amount; strcpy(newNode->description, description); while (first!=NULL) { if (strcmp(surname,first->surname)>0) second=first; else if (strcmp(surname,first->surname)==0) { if (strcmp(short_name,first->short_name)>0) second=first; } first=first->next; } if (second==NULL) { newNode->next=head; head=newNode; } else //SEGMENTATION APPEARS WHEN IT GETS HERE! { tmp=second->next; newNode->next=tmp; first->next=newNode; } } UPDATE: SegFault appears only when it gets on the 'else' loop of InsertSort() function. I observed that segmentation fault appears when i try to put on the list names that are after it. For example, if in the list exists: [Name:b Surname:b Amount:6 Description:b] [Name:c Surname:c Amount:5 Description:c] [Name:d Surname:d Amount:4 Description:d] [Name:e Surname:e Amount:3 Description:e] [Name:g Surname:g Amount:2 Description:g] [Name:x Surname:x Amount:1 Description:x] and i put: " x z 77 gege" there is a segmentation but if i put "x a 77 gege" it continues normally....

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  • How to make my robot move in a rectangular path along the black tape?

    - by Sahat
    I am working on a robot, it's part of the summer robotics workshop in our college. We are using C-STAMP micro controllers by A-WIT. I was able to make it move, turn left, turn right, move backward. I have even managed to make it go along the black tape using a contrast sensor. I send the robot at 30-45 degrees toward the black tape on the table and it aligns itself and starts to move along the black tape. It jerks a little, probably due to my programming logic below, it's running a while loop and constantly checking if statements, so it ends up trying to turn left and right every few milliseconds, which explains the jerking part. But it's okay, it works, not as smooth as I want it to work but it works! Problem is that I can't make my robot go into a rectangular path of the black tape. As soon as it reaches the corner it just keeps going straight instead of making a left/right turn. Here's my attempt. The following code is just part of the code. My 2 sensors are located right underneath the robot, next to the front wheel, almost at the floor level. It has "index" value ranging from 0 to 8. I believe it's 8 when you have a lot of light coming into the sensor , and 0 when it's nearly pitch black. So when the robot moves into the black-tape-zone, the index value drops, and based on that I have an if-statement telling my robot to either turn left or right. To avoid confusion I didn't post the entire source code, but only the logical part responsible for the movement of my robot along the black tape. while(1) { // don't worry about these. // 10 and 9 represent Sensor's PIN location on the motherboard V = ANALOGIN(10, 1, 0, 0, 0); V2 = ANALOGIN(9, 1, 0, 0, 0); // i got this "formula" from the example in my Manual. // V stands for voltage of the sensor. // it gives me the index value of the sensor. 0 = darkest, 8 = lightest. index = ((-(V - 5) / 5) * 8 + 0.5); index2 = ((-(V2 - 5) / 5) * 8 + 0.5); // i've tweaked the position of the sensors so index > 7 is just right number. // the robot will move anywhere on the table just fine with index > 7. // as soon as it drops to or below 7 (i.e. finds black tape), the robot will // either turn left or right and then go forward. // lp & rp represent left-wheel pin and right-wheel pin, 1 means run forever. // if i change it from 1 to 100, it will go forward for 100ms. if (index > 7 && index2 > 7) goForward(lp, rp, 1); if (index <= 7) { turnLeft(lp, rp, 1); goForward(lp, rp, 1); // this is the tricky part. i've added this code last minute // trying to make my robot turn, but i didn't work. if (index > 4) { turnLeft(lp, rp, 1); goForward(lp, rp, 1); } } else if (index2 <= 7) { turnRight(lp, rp, 1); goForward(lp, rp, 1); // this is also the last minute addition. it's same code as above // but it's for the 2nd sensor. if (index2 > 4) { turnRight(lp, rp, 1); goForward(lp, rp, 1); } } I've spent the entire day trying to figure it out. I've pretty much exhausted all avenues. Asking for the solution on stackoverflow is my very last option now. Thanks in advance! If you have any questions about the code, let me know, but comments should be self-explanatory.

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  • How do I name an array key with a key inside the array

    - by Confused
    I have some data, yes, data. This data came from a MySQL query and it will always contain 4 items, always. I want to cache that data in an array table for use later within a web page but I want to keep the keys from the query and separate out each grouping within a multidimensional array. However to save time iterating through the array each time I want to find a given group of data, I want to call the keys of the first array the same as the ID key which is always the first key within each four items. At the minute I'm using this code: function mysql_fetch_full_result_array($result) { $table_result=array(); $r=0; while($row = mysql_fetch_assoc($result)){ $arr_row=array(); $c=0; while ($c < mysql_num_fields($result)) { $col = mysql_fetch_field($result, $c); $arr_row[$col -> name] = $row[$col -> name]; $c++; } $table_result[$r] = $arr_row; $r++; } return $table_result; } I'm currently testing this using 3 unique users, so I'm getting three rows back from the query and the data from this function ends up in the format: [0]=> . . [id] => 1 . . [name] => random name . . [tel] => random tel . . [post] => post code data [1]=> . . [id] => 34 . . [name] => random name . . [tel] => random tel . . [post] => post code data [2]=> . . [id] => 56 . . [name] => random name . . [tel] => random tel . . [post] => post code data So how do I alter the code to instead of the keys [0], [1], [2] give me the output: [1]=> . . [id] => 1 . . [name] => random name . . [tel] => random tel . . [post] => post code data [34]=> . . [id] => 34 . . [name] => random name . . [tel] => random tel . . [post] => post code data [56]=> . . [id] => 56 . . [name] => random name . . [tel] => random tel . . [post] => post code data I don't mind if the main array keys are strings of numbers rather than numbers but I'm a bit stuck, I tried changing the $table_result[$r] = $arr_row; part to read $table_result[$result['id']] = $arr_row; but that just outputs an array of one person. I know I need another loop but I'm struggling to work out how to write it.

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  • libcurl - unable to download a file

    - by marmistrz
    I'm working on a program which will download lyrics from sites like AZLyrics. I'm using libcurl. It's my code lyricsDownloader.cpp #include "lyricsDownloader.h" #include <curl/curl.h> #include <cstring> #include <iostream> #define DEBUG 1 ///////////////////////////////////////////////////////////////////////////// size_t lyricsDownloader::write_data_to_var(char *ptr, size_t size, size_t nmemb, void *userdata) // this function is a static member function { ostringstream * stream = (ostringstream*) userdata; size_t count = size * nmemb; stream->write(ptr, count); return count; } string AZLyricsDownloader::toProviderCode() const { /*this creates an url*/ } CURLcode AZLyricsDownloader::download() { CURL * handle; CURLcode err; ostringstream buff; handle = curl_easy_init(); if (! handle) return static_cast<CURLcode>(-1); // set verbose if debug on curl_easy_setopt( handle, CURLOPT_VERBOSE, DEBUG ); curl_easy_setopt( handle, CURLOPT_URL, toProviderCode().c_str() ); // set the download url to the generated one curl_easy_setopt(handle, CURLOPT_WRITEDATA, &buff); curl_easy_setopt(handle, CURLOPT_WRITEFUNCTION, &AZLyricsDownloader::write_data_to_var); err = curl_easy_perform(handle); // The segfault should be somewhere here - after calling the function but before it ends cerr << "cleanup\n"; curl_easy_cleanup(handle); // copy the contents to text variable lyrics = buff.str(); return err; } main.cpp #include <QString> #include <QTextEdit> #include <iostream> #include "lyricsDownloader.h" int main(int argc, char *argv[]) { AZLyricsDownloader dl(argv[1], argv[2]); dl.perform(); QTextEdit qtexted(QString::fromStdString(dl.lyrics)); cout << qPrintable(qtexted.toPlainText()); return 0; } When running ./maelyrica Anthrax Madhouse I'm getting this logged from curl * About to connect() to azlyrics.com port 80 (#0) * Trying 174.142.163.250... * connected * Connected to azlyrics.com (174.142.163.250) port 80 (#0) > GET /lyrics/anthrax/madhouse.html HTTP/1.1 Host: azlyrics.com Accept: */* < HTTP/1.1 301 Moved Permanently < Server: nginx/1.0.12 < Date: Thu, 05 Jul 2012 16:59:21 GMT < Content-Type: text/html < Content-Length: 185 < Connection: keep-alive < Location: http://www.azlyrics.com/lyrics/anthrax/madhouse.html < Segmentation fault Strangely, the file is there. The same error is displayed when there's no such page (redirect to azlyrics.com mainpage) What am I doing wrong? Thanks in advance EDIT: I made the function for writing data static, but this changes nothing. Even wget seems to have problems $ wget http://www.azlyrics.com/lyrics/anthrax/madhouse.html --2012-07-06 10:36:05-- http://www.azlyrics.com/lyrics/anthrax/madhouse.html Resolving www.azlyrics.com... 174.142.163.250 Connecting to www.azlyrics.com|174.142.163.250|:80... connected. HTTP request sent, awaiting response... No data received. Retrying. Why does opening the page in a browser work and wget/curl not? EDIT2: After adding this: curl_easy_setopt(handle, CURLOPT_FOLLOWLOCATION, 1); The log is: * About to connect() to azlyrics.com port 80 (#0) * Trying 174.142.163.250... * connected * Connected to azlyrics.com (174.142.163.250) port 80 (#0) > GET /lyrics/anthrax/madhouse.html HTTP/1.1 Host: azlyrics.com Accept: */* < HTTP/1.1 301 Moved Permanently < Server: nginx/1.0.12 < Date: Fri, 06 Jul 2012 09:09:47 GMT < Content-Type: text/html < Content-Length: 185 < Connection: keep-alive < Location: http://www.azlyrics.com/lyrics/anthrax/madhouse.html < * Ignoring the response-body * Connection #0 to host azlyrics.com left intact * Issue another request to this URL: 'http://www.azlyrics.com/lyrics/anthrax/madhouse.html' * About to connect() to www.azlyrics.com port 80 (#1) * Trying 174.142.163.250... * connected * Connected to www.azlyrics.com (174.142.163.250) port 80 (#1) > GET /lyrics/anthrax/madhouse.html HTTP/1.1 Host: www.azlyrics.com Accept: */* < HTTP/1.1 200 OK < Server: nginx/1.0.12 < Date: Fri, 06 Jul 2012 09:09:47 GMT < Content-Type: text/html < Transfer-Encoding: chunked < Connection: keep-alive < Segmentation fault

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  • Watching setTimeout loops so that only one is running at a time.

    - by DA
    I'm creating a content rotator in jQuery. 5 items total. Item 1 fades in, pauses 10 seconds, fades out, then item 2 fades in. Repeat. Simple enough. Using setTimeout I can call a set of functions that create a loop and will repeat the process indefinitely. I now want to add the ability to interrupt this rotator at any time by clicking on a navigation element to jump directly to one of the content items. I originally started going down the path of pinging a variable constantly (say every half second) that would check to see if a navigation element was clicked and, if so, abandon the loop, then restart the loop based on the item that was clicked. The challenge I ran into was how to actually ping a variable via a timer. The solution is to dive into JavaScript closures...which are a little over my head but definitely something I need to delve into more. However, in the process of that, I came up with an alternative option that actually seems to be better performance-wise (theoretically, at least). I have a sample running here: http://jsbin.com/uxupi/14 (It's using console.log so have fireBug running) Sample script: $(document).ready(function(){ var loopCount = 0; $('p#hello').click(function(){ loopCount++; doThatThing(loopCount); }) function doThatOtherThing(currentLoopCount) { console.log('doThatOtherThing-'+currentLoopCount); if(currentLoopCount==loopCount){ setTimeout(function(){doThatThing(currentLoopCount)},5000) } } function doThatThing(currentLoopCount) { console.log('doThatThing-'+currentLoopCount); if(currentLoopCount==loopCount){ setTimeout(function(){doThatOtherThing(currentLoopCount)},5000); } } }) The logic being that every click of the trigger element will kick off the loop passing into itself a variable equal to the current value of the global variable. That variable gets passed back and forth between the functions in the loop. Each click of the trigger also increments the global variable so that subsequent calls of the loop have a unique local variable. Then, within the loop, before the next step of each loop is called, it checks to see if the variable it has still matches the global variable. If not, it knows that a new loop has already been activated so it just ends the existing loop. Thoughts on this? Valid solution? Better options? Caveats? Dangers? UPDATE: I'm using John's suggestion below via the clearTimeout option. However, I can't quite get it to work. The logic is as such: var slideNumber = 0; var timeout = null; function startLoop(slideNumber) { ...do stuff here to set up the slide based on slideNumber... slideFadeIn() } function continueCheck(){ if (timeout != null) { // cancel the scheduled task. clearTimeout(timeout); timeout = null; return false; }else{ return true; } }; function slideFadeIn() { if (continueCheck){ // a new loop hasn't been called yet so proceed... // fade in the LI $currentListItem.fadeIn(fade, function() { if(multipleFeatures){ timeout = setTimeout(slideFadeOut,display); } }); }; function slideFadeOut() { if (continueLoop){ // a new loop hasn't been called yet so proceed... slideNumber=slideNumber+1; if(slideNumber==features.length) { slideNumber = 0; }; timeout = setTimeout(function(){startLoop(slideNumber)},100); }; startLoop(slideNumber); The above kicks of the looping. I then have navigation items that, when clicked, I want the above loop to stop, then restart with a new beginning slide: $(myNav).click(function(){ clearTimeout(timeout); timeout = null; startLoop(thisItem); }) If I comment out 'startLoop...' from the click event, it, indeed, stops the initial loop. However, if I leave that last line in, it doesn't actually stop the initial loop. Why? What happens is that both loops seem to run in parallel for a period. So, when I click my navigation, clearTimeout is called, which clears it.

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  • How do I prevent TCP connection freezes over an OpenVPN network?

    - by Jason R
    New details added at the end of this question; it's possible that I'm zeroing in on the cause. I have a UDP OpenVPN-based VPN set up in tap mode (I need tap because I need the VPN to pass multicast packets, which doesn't seem to be possible with tun networks) with a handful of clients across the Internet. I've been experiencing frequent TCP connection freezes over the VPN. That is, I will establish a TCP connection (e.g. an SSH connection, but other protocols have similar issues), and at some point during the session, it seems that traffic will cease being transmitted over that TCP session. This seems to be related to points at which large data transfers occur, such as if I execute an ls command in an SSH session, or if I cat a long log file. Some Google searches turn up a number of answers like this previous one on Server Fault, indicating that the likely culprit is an MTU issue: that during periods of high traffic, the VPN is trying to send packets that get dropped somewhere in the pipes between the VPN endpoints. The above-linked answer suggests using the following OpenVPN configuration settings to mitigate the problem: fragment 1400 mssfix This should limit the MTU used on the VPN to 1400 bytes and fix the TCP maximum segment size to prevent the generation of any packets larger than that. This seems to mitigate the problem a bit, but I still frequently see the freezes. I've tried a number of sizes as arguments to the fragment directive: 1200, 1000, 576, all with similar results. I can't think of any strange network topology between the two ends that could trigger such a problem: the VPN server is running on a pfSense machine connected directly to the Internet, and my client is also connected directly to the Internet at another location. One other strange piece of the puzzle: if I run the tracepath utility, then that seems to band-aid the problem. A sample run looks like: [~]$ tracepath -n 192.168.100.91 1: 192.168.100.90 0.039ms pmtu 1500 1: 192.168.100.91 40.823ms reached 1: 192.168.100.91 19.846ms reached Resume: pmtu 1500 hops 1 back 64 The above run is between two clients on the VPN: I initiated the trace from 192.168.100.90 to the destination of 192.168.100.91. Both clients were configured with fragment 1200; mssfix; in an attempt to limit the MTU used on the link. The above results would seem to suggest that tracepath was able to detect a path MTU of 1500 bytes between the two clients. I would assume that it would be somewhat smaller due to the fragmentation settings specified in the OpenVPN configuration. I found that result somewhat strange. Even stranger, however: if I have a TCP connection in the stalled state (e.g. an SSH session with a directory listing that froze in the middle), then executing the tracepath command shown above causes the connection to start up again! I can't figure out any reasonable explanation for why this would be the case, but I feel like this might be pointing toward a solution to ultimately eradicate the problem. Does anyone have any recommendations for other things to try? Edit: I've come back and looked at this a bit further, and have found only more confounding information: I set the OpenVPN connection to fragment at 1400 bytes, as shown above. Then, I connected to the VPN from across the Internet and used Wireshark to look at the UDP packets that were sent to the VPN server while the stall occurred. None were greater than the specified 1400 byte count, so the fragmentation seems to be functioning properly. To verify that even a 1400-byte MTU would be sufficient, I pinged the VPN server using the following (Linux) command: ping <host> -s 1450 -M do This (I believe) sends a 1450-byte packet with fragmentation disabled (I at least verified that it didn't work if I set it to an obviously-too-large value like 1600 bytes). These seem to work just fine; I get replies back from the host with no issue. So, maybe this isn't an MTU issue at all. I'm just confused as to what else it might be! Edit 2: The rabbit hole just keeps getting deeper: I've now isolated the problem a bit more. It seems to be related to the exact OS that the VPN client uses. I have successfully duplicated the problem on at least three Ubuntu machines (versions 12.04 through 13.04). I can reliably duplicate an SSH connection freeze within a minute or so by just cat-ing a large log file. However, if I do the same test using a CentOS 6 machine as a client, then I don't see the problem! I've tested using the exact same OpenVPN client version as I was using on the Ubuntu machines. I can cat log files for hours without seeing the connection freeze. This seems to provide some insight as to the ultimate cause, but I'm just not sure what that insight is. I have examined the traffic over the VPN using Wireshark. I'm not a TCP expert, so I'm not sure what to make of the gory details, but the gist is that at some point, a UDP packet gets dropped due to the limited bandwidth of the Internet link, causing TCP retransmissions inside the VPN tunnel. On the CentOS client, these retransmissions occur properly and things move on happily. At some point with the Ubuntu clients, though, the remote end starts retransmitting the same TCP segment over and over (with the transmit delay increasing between each retransmission). The client sends what looks like a valid TCP ACK to each retransmission, but the remote end still continues to transmit the same TCP segment periodically. This extends ad infinitum and the connection stalls. My question here would be: Does anyone have any recommendations for how to troubleshoot and/or determine the root cause of the TCP issue? It's as if the remote end isn't accepting the ACK messages sent by the VPN client. One common difference between the CentOS node and the various Ubuntu releases is that Ubuntu has a much more recent Linux kernel version (from 3.2 in Ubuntu 12.04 to 3.8 in 13.04). A pointer to some new kernel bug maybe? I'm assuming that if that were so, then I wouldn't be the only one experiencing the problem; I don't think this seems like a particularly exotic setup.

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  • Elapsed time of running a C program

    - by yCalleecharan
    Hi, I would like to know what lines of C code to add to a program so that it tells me the total time that the program takes to run. I guess there should be counter initialization near the beginning of main and one after the main function ends. Is the right header clock.h? Thanks a lot... Update I have a Win Xp machine. Is it just adding clock() at the beginning and another clock() at the end of the program? Then I can estimate the time difference. Yes, you're right it's time.h. Here's my code: #include <stdio.h> #include <stdlib.h> #include <math.h> #include <share.h> #include <time.h> void f(long double fb[], long double fA, long double fB); int main() { clock_t start, end; start = clock(); const int ARRAY_SIZE = 11; long double* z = (long double*) malloc(sizeof (long double) * ARRAY_SIZE); int i; long double A, B; if (z == NULL) { printf("Out of memory\n"); exit(-1); } A = 0.5; B = 2; for (i = 0; i < ARRAY_SIZE; i++) { z[i] = 0; } z[1] = 5; f(z, A, B); for (i = 0; i < ARRAY_SIZE; i++) printf("z is %.16Le\n", z[i]); free(z); z = NULL; end = clock(); printf("Took %ld ticks\n", end-start); printf("Took %f seconds\n", (double)(end-start)/CLOCKS_PER_SEC); return 0; } void f(long double fb[], long double fA, long double fB) { fb[0] = fb[1]* fA; fb[1] = fb[1] - 1; return; } Some errors with MVS2008: testim.c(16) : error C2143: syntax error : missing ';' before 'const' testim.c(18) :error C2143: syntax error : missing ';' before 'type' testim.c(20) :error C2143: syntax error : missing ';' before 'type' testim.c(21) :error C2143: syntax error : missing ';' before 'type' testim.c(23) :error C2065: 'z' : undeclared identifier testim.c(23) :warning C4047: '==' : 'int' differs in levels of indirection from 'void *' testim.c(28) : error C2065: 'A' : undeclared identifier testim.c(28) : warning C4244: '=' : conversion from 'double' to 'int', possible loss of data and it goes to 28 errors. Note that I don't have any errors/warnings without your clock codes. LATEST NEWS: I unfortunately didn't get a good reply here. But after a search on Google, the code is working. Here it is: #include <stdio.h> #include <stdlib.h> #include <math.h> #include <share.h> #include <time.h> void f(long double fb[], long double fA, long double fB); int main() { clock_t start = clock(); const int ARRAY_SIZE = 11; long double* z = (long double*) malloc(sizeof (long double) * ARRAY_SIZE); int i; long double A, B; if (z == NULL) { printf("Out of memory\n"); exit(-1); } A = 0.5; B = 2; for (i = 0; i < ARRAY_SIZE; i++) { z[i] = 0; } z[1] = 5; f(z, A, B); for (i = 0; i < ARRAY_SIZE; i++) printf("z is %.16Le\n", z[i]); free(z); z = NULL; printf("Took %f seconds\n", ((double)clock()-start)/CLOCKS_PER_SEC); return 0; } void f(long double fb[], long double fA, long double fB) { fb[0] = fb[1]* fA; fb[1] = fb[1] - 1; return; } Cheers

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  • OpenGL + cgFX Alpha Blending failure

    - by dopplex
    I have a shader that needs to additively blend to its output render target. While it had been fully implemented and working, I recently refactored and have done something that is causing the alpha blending to not work anymore. I'm pretty sure that the problem is somewhere in my calls to either OpenGL or cgfx - but I'm currently at a loss for where exactly the problem is, as everything looks like it is set up properly for alpha blending to occur. No OpenGL or cg framework errors are showing up, either. For some context, what I'm doing here is taking a buffer which contains screen position and luminance values for each pixel, copying it to a PBO, and using it as the vertex buffer for drawing GL_POINTS. Everything except for the alpha blending appears to be working as expected. I've confirmed both that the input vertex buffer has the correct values, and that my vertex and fragment shaders are outputting the points to the correct locations and with the correct luminance values. The way that I've arrived at the conclusion that the Alpha blending was broken is by making my vertex shader output every point to the same screen location and then setting the pixel shader to always output a value of float4(0.5) for that pixel. Invariably, the end color (dumped afterwards) ends up being float4(0.5). The confusing part is that as far as I can tell, everything is properly set for alpha blending to occur. The cgfx pass has the two following state assignments (among others - I'll put a full listing at the end): BlendEnable = true; BlendFunc = int2(One, One); This ought to be enough, since I am calling cgSetPassState() - and indeed, when I use glGets to check the values of GL_BLEND_SRC, GL_BLEND_DEST, GL_BLEND, and GL_BLEND_EQUATION they all look appropriate (GL_ONE, GL_ONE, GL_TRUE, and GL_FUNC_ADD). This check was done immediately after the draw call. I've been looking around to see if there's anything other than blending being enabled and the blending function being correctly set that would cause alpha blending not to occur, but without any luck. I considered that I could be doing something wrong with GL, but GL is telling me that blending is enabled. I doubt it's cgFX related (as otherwise the GL state wouldn't even be thinking it was enabled) but it still fails if I explicitly use GL calls to set the blend mode and enable it. Here's the trimmed down code for starting the cgfx pass and the draw call: CGtechnique renderTechnique = Filter->curTechnique; TEXUNITCHECK; CGpass pass = cgGetFirstPass(renderTechnique); TEXUNITCHECK; while (pass) { cgSetPassState(pass); cgUpdatePassParameters(pass); //drawFSPointQuadBuff((void*)PointQuad); drawFSPointQuadBuff((void*)LumPointBuffer); TEXUNITCHECK; cgResetPassState(pass); pass = cgGetNextPass(pass); }; and the function with the draw call: void drawFSPointQuadBuff(void* args) { PointBuffer* pointBuffer = (PointBuffer*)args; FBOERRCHECK; glClear(GL_COLOR_BUFFER_BIT); GLERRCHECK; glPointSize(1.0); GLERRCHECK; glEnableClientState(GL_VERTEX_ARRAY); GLERRCHECK; glEnable(GL_POINT_SMOOTH); if (pointBuffer-BufferObject) { glBindBufferARB(GL_ARRAY_BUFFER_ARB, (unsigned int)pointBuffer-BufData); glVertexPointer(pointBuffer-numComp, GL_FLOAT, 0, 0); } else { glVertexPointer(pointBuffer-numComp, GL_FLOAT, 0, pointBuffer-BufData); }; GLERRCHECK; glDrawArrays(GL_POINTS, 0, pointBuffer-numElem); GLboolean testBool; glGetBooleanv(GL_BLEND, &testBool); int iblendColor, iblendDest, iblendEquation, iblendSrc; glGetIntegerv(GL_BLEND_SRC, &iblendSrc); glGetIntegerv(GL_BLEND_DST, &iblendDest); glGetIntegerv(GL_BLEND_EQUATION, &iblendEquation); if (iblendEquation == GL_FUNC_ADD) { cerr << "Correct func" << endl; }; GLERRCHECK; if (pointBuffer-BufferObject) { glBindBufferARB(GL_ARRAY_BUFFER_ARB,0); } GLERRCHECK; glDisableClientState(GL_VERTEX_ARRAY); GLERRCHECK; }; Finally, here is the full state setting of the shader: AlphaTestEnable = false; DepthTestEnable = false; DepthMask = false; ColorMask = true; CullFaceEnable = false; BlendEnable = true; BlendFunc = int2(One, One); FragmentProgram = compile glslf std_PS(); VertexProgram = compile glslv bilatGridVS2();

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  • Something like a manual refresh is needed angularjs, and a $digest() iterations error

    - by Tony Ennis
    (post edited again, new comments follow this line) I'm changing the title of this posting since it was misleading - I was trying to fix a symptom. I was unable to figure out why the code was breaking with a $digest() iterations error. A plunk of my code worked fine. I was totally stuck, so I decided to make my code a little more Angular-like. One anti-pattern I had implemented was to hide my model behind my controller by adding getters/setters to the controller. I tore all that out and instead put the model into the $scope since I had read that was proper Angular. To my surprise, the $digest() iterations error went away. I do not exactly know why and I do not have the intestinal fortitude to put the old code back and figure it out. I surmise that by involving the controller in the get/put of the data I added a dependency under the hood. I do not understand it. edit #2 ends here. (post edited, see EDIT below) I was working through my first Error: 10 $digest() iterations reached. Aborting! error today. I solved it this way: <div ng-init="lineItems = ctrl.getLineItems()"> <tr ng-repeat="r in lineItems"> <td>{{r.text}}</td> <td>...</td> <td>{{r.price | currency}}</td> </tr </div> Now a new issue has arisen - the line items I'm producing can be modified by another control on the page. It's a text box for a promo code. The promo code adds a discount to the lineItem array. It would show up if I could ng-repeat over ctrl.getLineItems(). Since the ng-repeat is looking at a static variable, not the actual model, it doesn't see that the real line items have changed and thus the promotional discount doesn't get displayed until I refresh the browser. Here's the HTML for the promo code: <input type="text" name="promo" ng-model="ctrl.promoCode"/> <button ng-click="ctrl.applyPromoCode()">apply promo code</button> The input tag is writing the value to the model. The bg-click in the button is invoking a function that will apply the code. This could change the data behind the lineItems. I have been advised to use $scope.apply(...). However, since this is applied as a matter of course by ng-click is isn't going to do anything. Indeed, if I add it to ctrl.applyPromoCode(), I get an error since an .apply() is already in progress. I'm at a loss. EDIT The issue above is probably the result of me fixing of symptom, not a problem. Here is the original HTML that was dying with the 10 $digest() iterations error. <table> <tr ng-repeat="r in ctrl.getLineItems()"> <td>{{r.text}}</td> <td>...</td> <td>{{r.price | currency}}</td> </tr> </table> The ctrl.getLineItems() function doesn't do much but invoke a model. I decided to keep the model out of the HTML as much as I could. this.getLineItems = function() { var total = 0; this.lineItems = []; this.lineItems.push({text:"Your quilt will be "+sizes[this.size].block_size+" squares", price:sizes[this.size].price}); total = sizes[this.size].price; this.lineItems.push({text: threads[this.thread].narrative, price:threads[this.thread].price}); total = total + threads[this.thread].price; if (this.sashing) { this.lineItems.push({text:"Add sashing", price: this.getSashingPrice()}); total = total + sizes[this.size].sashing; } else { this.lineItems.push({text:"No sashing", price:0}); } if(isNaN(this.promo)) { this.lineItems.push({text:"No promo code", price:0}); } else { this.lineItems.push({text:"Promo code", price: promos[this.promo].price}); total = total + promos[this.promo].price; } this.lineItems.push({text:"Shipping", price:this.shipping}); total = total + this.shipping; this.lineItems.push({text:"Order Total", price:total}); return this.lineItems; }; And the model code assembled an array of objects based upon the items selected. I'll abbreviate the class as it croaks as long as the array has a row. function OrderModel() { this.lineItems = []; // Result of the lineItems call ... this.getLineItems = function() { var total = 0; this.lineItems = []; ... this.lineItems.push({text:"Order Total", price:total}); return this.lineItems; }; }

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  • Saving a Join Model

    - by Thorpe Obazee
    I've been reading the cookbook for a while now and still don't get how I'm supposed to do this: My original problem was this: A related Model isn't being validated From RabidFire's commment: If you want to count the number of Category models that a new Post is associated with (on save), then you need to do this in the beforeSave function as I've mentioned. As you've currently set up your models, you don't need to use the multiple rule anywhere. If you really, really want to validate against a list of Category IDs for some reason, then create a join model, and validate category_id with the multiple rule there. Now, I have these models and are now validating. The problem now is that data isn't being saved in the Join Table: class Post extends AppModel { var $name = 'Post'; var $hasMany = array( 'CategoryPost' => array( 'className' => 'CategoryPost' ) ); var $belongsTo = array( 'Page' => array( 'className' => 'Page' ) ); class Category extends AppModel { var $name = 'Category'; var $hasMany = array( 'CategoryPost' => array( 'className' => 'CategoryPost' ) ); class CategoryPost extends AppModel { var $name = 'CategoryPost'; var $validate = array( 'category_id' => array( 'rule' => array('multiple', array('in' => array(1, 2, 3, 4))), 'required' => FALSE, 'message' => 'Please select one, two or three options' ) ); var $belongsTo = array( 'Post' => array( 'className' => 'Post' ), 'Category' => array( 'className' => 'Category' ) ); This is the new Form: <div id="content-wrap"> <div id="main"> <h2>Add Post</h2> <?php echo $this->Session->flash();?> <div> <?php echo $this->Form->create('Post'); echo $this->Form->input('Post.title'); echo $this->Form->input('CategoryPost.category_id', array('multiple' => 'checkbox')); echo $this->Form->input('Post.body', array('rows' => '3')); echo $this->Form->input('Page.meta_keywords'); echo $this->Form->input('Page.meta_description'); echo $this->Form->end('Save Post'); ?> </div> <!-- main ends --> </div> The data I am producing from the form is as follows: Array ( [Post] => Array ( [title] => 1234 [body] => 1234 ) [CategoryPost] => Array ( [category_id] => Array ( [0] => 1 [1] => 2 ) ) [Page] => Array ( [meta_keywords] => 1234 [meta_description] => 1234 [title] => 1234 [layout] => index ) ) UPDATE: controller action //Controller action function admin_add() { // pr(Debugger::trace()); $this->set('categories', $this->Post->CategoryPost->Category->find('list')); if ( ! empty($this->data)) { $this->data['Page']['title'] = $this->data['Post']['title']; $this->data['Page']['layout'] = 'index'; debug($this->data); if ($this->Post->saveAll($this->data)) { $this->Session->setFlash('Your post has been saved', 'flash_good'); $this->redirect($this->here); } } } UPDATE #2: Should I just do this manually? The problem is that the join tables doesn't have things saved in it. Is there something I'm missing? UPDATE #3 RabidFire gave me a solution. I already did this before and am quite surprised as so why it didn't work. Thus, me asking here. The reason I think there is something wrong. I don't know where: Post beforeSave: function beforeSave() { if (empty($this->id)) { $this->data[$this->name]['uri'] = $this->getUniqueUrl($this->data[$this->name]['title']); } if (isset($this->data['CategoryPost']['category_id']) && is_array($this->data['CategoryPost']['category_id'])) { echo 'test'; $categoryPosts = array(); foreach ($this->data['CategoryPost']['category_id'] as $categoryId) { $categoryPost = array( 'category_id' => $categoryId ); array_push($categoryPosts, $categoryPost); } $this->data['CategoryPost'] = $categoryPosts; } debug($this->data); // Gives RabidFire's correct array for saving. return true; } My Post action: function admin_add() { // pr(Debugger::trace()); $this->set('categories', $this->Post->CategoryPost->Category->find('list')); if ( ! empty($this->data)) { $this->data['Page']['title'] = $this->data['Post']['title']; $this->data['Page']['layout'] = 'index'; debug($this->data); // First debug is giving the correct array as above. if ($this->Post->saveAll($this->data)) { debug($this->data); // STILL gives the above array. which shouldn't be because of the beforeSave in the Post Model // $this->Session->setFlash('Your post has been saved', 'flash_good'); // $this->redirect($this->here); } } }

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  • I asked this yesterday, after the input given I'm still having trouble implementing..

    - by Josh
    I'm not sure how to fix this or what I did wrong, but whenever I enter in a value it just closes out the run prompt. So, seems I do have a problem somewhere in my coding. Whenever I run the program and input a variable, it always returns the same answer.."The content at location 76 is 0." On that note, someone told me that "I don't know, but I suspect that Program A incorrectly has a fixed address being branched to on instructions 10 and 11." - mctylr but I'm not sure how to fix that.. I'm trying to figure out how to incorporate this idea from R Samuel Klatchko.. I'm still not sure what I'm missing but I can't get it to work.. const int OP_LOAD = 3; const int OP_STORE = 4; const int OP_ADD = 5; ... const int OP_LOCATION_MULTIPLIER = 100; mem[0] = OP_LOAD * OP_LOCATION_MULTIPLIER + ...; mem[1] = OP_ADD * OP_LOCATION_MULTIPLIER + ...; operand = memory[ j ] % OP_LOCATION_MULTIPLIER; operation = memory[ j ] / OP_LOCATION_MULTIPLIER; I'm new to programming, I'm not the best, so I'm going for simplicity. Also this is an SML program. Anyway, this IS a homework assignment and I'm wanting a good grade on this. So I was looking for input and making sure this program will do what I'm hoping they are looking for. Anyway, here are the instructions: Write SML (Simpletron Machine language) programs to accomplish each of the following task: A) Use a sentinel-controlled loop to read positive number s and compute and print their sum. Terminate input when a neg number is entered. B) Use a counter-controlled loop to read seven numbers, some positive and some negative, and compute + print the avg. C) Read a series of numbers, and determine and print the largest number. The first number read indicates how many numbers should be processed. Without further a due, here is my program. All together. int main() { const int READ = 10; const int WRITE = 11; const int LOAD = 20; const int STORE = 21; const int ADD = 30; const int SUBTRACT = 31; const int DIVIDE = 32; const int MULTIPLY = 33; const int BRANCH = 40; const int BRANCHNEG = 41; const int BRANCHZERO = 41; const int HALT = 43; int mem[100] = {0}; //Making it 100, since simpletron contains a 100 word mem. int operation; //taking the rest of these variables straight out of the book seeing as how they were italisized. int operand; int accum = 0; // the special register is starting at 0 int j; // This is for part a, it will take in positive variables in a sent-controlled loop and compute + print their sum. Variables from example in text. memory [0] = 1010; memory [01] = 2009; memory [02] = 3008; memory [03] = 2109; memory [04] = 1109; memory [05] = 4300; memory [06] = 1009; j = 0; //Makes the variable j start at 0. while ( true ) { operand = memory[ j ]%100; // Finds the op codes from the limit on the memory (100) operation = memory[ j ]/100; //using a switch loop to set up the loops for the cases switch ( operation ){ case 10: //reads a variable into a word from loc. Enter in -1 to exit cout <<"\n Input a positive variable: "; cin >> memory[ operand ]; break; case 11: // takes a word from location cout << "\n\nThe content at location " << operand << "is " << memory[operand]; break; case 20:// loads accum = memory[ operand ]; break; case 21: //stores memory[ operand ] = accum; break; case 30: //adds accum += mem[operand]; break; case 31: // subtracts accum-= memory[ operand ]; break; case 32: //divides accum /=(memory[ operand ]); break; case 33: // multiplies accum*= memory [ operand ]; break; case 40: // Branches to location j = -1; break; case 41: //branches if acc. is < 0 if (accum < 0) j = 5; break; case 42: //branches if acc = 0 if (accum == 0) j = 5; break; case 43: // Program ends exit(0); break; } j++; } return 0; }

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  • C -Segmentation fault after the 4th call of the function!

    - by FILIaS
    It seems at least weird to me... The program runs normally.But after I call the enter() function for the 4th time,there is a segmentation fault!I would appreciate any help. With the following function enter() I wanna add user commands' datas to a list. [Some part of the code is already posted on another question of me, but I think I should post it again...as it's a different problem I'm facing now.] /* struct for all the datas that user enters on file*/ typedef struct catalog { char short_name[50]; char surname[50]; signed int amount; char description[1000]; struct catalog *next; }catalog,*catalogPointer; catalogPointer current; catalogPointer head = NULL; void enter(void) //user command: i <name> <surname> <amount> <description> { int n,j=2,k=0; char temp[1500]; char *short_name,*surname,*description; signed int amount; char* params = strchr(command,' ') + 1; //strchr returns a pointer to the 1st space on the command.U want a pointer to the char right after that space. strcpy(temp, params); //params is saved as temp. char *curToken = strtok(temp," "); //strtok cuts 'temp' into strings between the spaces and saves them to 'curToken' printf("temp is:%s \n",temp); printf("\nWhat you entered for saving:\n"); for (n = 0; curToken; ++n) //until curToken ends: { if (curToken) { short_name = malloc(strlen(curToken) + 1); strncpy(short_name, curToken, sizeof (short_name)); } printf("Short Name: %s \n",short_name); curToken = strtok(NULL," "); if (curToken) { surname = malloc(strlen(curToken) + 1); strncpy(surname, curToken,sizeof (surname)); } printf("SurName: %s \n",surname); curToken = strtok(NULL," "); if (curToken) { //int * amount= malloc(sizeof (signed int *)); char *chk; amount = (int) strtol(curToken, &chk, 10); if (!isspace(*chk) && *chk != 0) fprintf(stderr,"Warning: expected integer value for amount, received %s instead\n",curToken); } printf("Amount: %d \n",amount); curToken = strtok(NULL,"\0"); if (curToken) { description = malloc(strlen(curToken) + 1); strncpy(description, curToken, sizeof (description)); } printf("Description: %s \n",description); break; } if (findEntryExists(head, surname,short_name) != NULL) //call function in order to see if entry exists already on the catalog printf("\nAn entry for <%s %s> is already in the catalog!\nNew entry not entered.\n",short_name,surname); else { printf("\nTry to entry <%s %s %d %s> in the catalog list!\n",short_name,surname,amount,description); newEntry(&head,short_name,surname,amount,description); printf("\n**Entry done!**\n"); } // Maintain the list in alphabetical order by surname. } catalogPointer findEntryExists (catalogPointer head, char num[],char first[]) { catalogPointer p = head; while (p != NULL && strcmp(p->surname, num) != 0 && strcmp(p->short_name,first) != 0) { p = p->next; } return p; } catalogPointer newEntry (catalog** headRef,char short_name[], char surname[], signed int amount, char description[]) { catalogPointer newNode = (catalogPointer)malloc(sizeof(catalog)); catalogPointer first; catalogPointer second; catalogPointer tmp; first=head; second=NULL; strcpy(newNode->short_name, short_name); strcpy(newNode->surname, surname); newNode->amount=amount; strcpy(newNode->description, description); //SEGMENTATION ON THE 4TH RUN OF PROGRAM STOPS HERE depending on the names each time it gets! while (first!=NULL) { if (strcmp(surname,first->surname)>0) second=first; else if (strcmp(surname,first->surname)==0) { if (strcmp(short_name,first->short_name)>0) second=first; } first=first->next; } if (second==NULL) { newNode->next=head; head=newNode; } else { tmp=second->next; newNode->next=tmp; first->next=newNode; } }

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  • jQuery doesn't work after an Ajax post

    - by user1758979
    I'm using jQuery to sort a list of entries, between <LI></LI> tags, and then an Ajax post to validate the order and 'update' the page with the content returned. $.ajax({url: "./test.php?id=<?php echo $id; ?>&action=modify", contenttype: "application/x-www-form-urlencoded;charset=utf-8", data: {myJson: data}, type: 'post', success: function(data) { $('html').html(data); OnloadFunction (); } }); Then, I lose the ability to sort the list (I'm not sure if clear...). I tried to move the content of the $(document).ready inside the OnloadFunction (), and call it with <script>OnloadFunction ();</script> inside the block dealing with the modifications to do : $action= $_GET['action']; if ($action == "modify") { // Code here } but it doesn't work... I can't figure out how to do that. Could anyone help ? I stripped out the main part of the code to keep only the essential (filename: test.php) <!DOCTYPE html PUBLIC "-//W3C//DTD XHTML 1.0 Transitional//EN" "http://www.w3.org/TR/xhtml1/DTD/xhtml1-transitional.dtd"> <html xmlns="http://www.w3.org/1999/xhtml"> <head> <meta http-equiv="Content-Type" content="text/html; charset=ISO-8859-1"> <script type="text/javascript" src="jquery-1.8.2.min.js"></script> <script type="text/javascript" src="jquery-ui-1.9.0.custom.min.js"></script> <script> $(document).ready(function(){ //alert("I am ready"); OnloadFunction (); }); function OnloadFunction () { $(function() { $("#SortColumn ul").sortable({ opacity: 0.6, cursor: 'move', update: function() {} }); }); //alert('OnloadFunction ends'); } function valider(){ var SortedId = new Array(); SortIdNb = 0; $('#SortColumn ul li').each(function() { SortedId.push(this.id); }); var data = { /* Real code contains an array with the <li> id */ CheckedId: "CheckedId", SortedId: SortedId, }; data = JSON.stringify(data); $.ajax({url: "./test.php?id=<?php echo $id; ?>&action=modify", contenttype: "application/x-www-form-urlencoded;charset=utf-8", data: {myJson: data}, type: 'post', success: function(data) { //alert(data); $('html').html(data); OnloadFunction (); } }); } </script> </head> <body> <? $action= $_GET['action']; $id = $_GET['id']; if ($id == 0) {$id=1;} $id += 1; if ($action == "modify") { echo "action: modify<br>"; echo "id (àvèc aççént$): ".$id."<br>"; // "(àvèc aççént$)" to check characters because character set is incorrect after the ajax post $data = json_decode($_POST['myJson'], true); // PHP code here to treat the new list send via the post and update the database print_r($data); } ?> <!-- PHP code here to get the following list from the database --> <div id="SortColumn"> <ul> <li id="recordsArray_1">recordsArray_1</li> <li id="recordsArray_2">recordsArray_2</li> <li id="recordsArray_3">recordsArray_3</li> <li id="recordsArray_4">recordsArray_4</li> <li id="recordsArray_5">recordsArray_5</li> </ul> </div> <input type="button" value="Modifier" onclick="valider();"> </body> </html>

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  • How to obtain listview information without refreshing the page?

    - by user1808098
    I am currently developing an Android Application for my Final Year Project. But to be honest I do not have any basic knowledges and everything started from scratch and referring to online tutorials a lot. Here is my question, I was trying to retrieve data from listview activity. There are two listview in my page using button. I was able to display the first listview but when it get data for the second listview, the data for first listview is disappeared because the page is refreshed, vice versa. What code should I modified to get both the data in the page? (Database not implemented yet) Please help, thanks a lot. Below are my codings. Codings for XML. <!-- Location --> <TextView android:id="@+id/TextViewLocation" android:layout_width="fill_parent" android:layout_height="wrap_content" android:layout_marginTop="5dip" android:layout_marginBottom="10dip" android:text="Location Information" android:gravity="center" android:textSize="15dip" android:textColor="#025f7c"/> <!-- Condition Label --> <TextView android:layout_width="fill_parent" android:layout_height="wrap_content" android:textColor="#372c24" android:text="Traffic Condition"/> <Button android:id="@+id/inputListView" android:layout_width="fill_parent" android:layout_height="wrap_content" android:layout_marginTop="5dip" android:layout_marginBottom="10dip" android:text="choose one..."/> <!-- Comment Label --> <TextView android:layout_width="fill_parent" android:layout_height="wrap_content" android:textColor="#372c24" android:text="What's Happening?"/> <Button android:id="@+id/inputListView2" android:layout_width="fill_parent" android:layout_height="wrap_content" android:layout_marginTop="5dip" android:layout_marginBottom="10dip" android:text="choose one..."/> <!-- Suggestion Label --> <TextView android:layout_width="fill_parent" android:layout_height="wrap_content" android:textColor="#372c24" android:text="Comments / Suggestion"/> <EditText android:layout_width="fill_parent" android:layout_height="80dp" android:layout_marginTop="5dip" android:layout_marginBottom="10dip" android:singleLine="true"/> <!-- Image button --> <Button android:id="@+id/btnImage" android:layout_width="fill_parent" android:layout_height="wrap_content" android:layout_marginTop="5dip" android:text="Upload Image"/> <!-- Report button --> <Button android:id="@+id/btnReportCheckin" android:layout_width="fill_parent" android:layout_height="wrap_content" android:layout_marginTop="5dip" android:text="Report"/> <!-- Link to Logout --> <TextView android:id="@+id/linkLogout" android:layout_width="fill_parent" android:layout_height="wrap_content" android:layout_marginTop="5dip" android:layout_marginBottom="40dip" android:text="Log Out" android:gravity="center" android:textSize="20dip" android:textColor="#025f7c"/> </LinearLayout> <!-- Check or Report Form Ends --> Codings for Activity Class public class CheckinActivity extends Activity { @Override public void onCreate(Bundle savedInstanceState) { super.onCreate(savedInstanceState); // Set View to checkin.xml setContentView(R.layout.checkin); /* TextView LocationView = (TextView) findViewById(R.id.TextViewLocation); Intent h = getIntent(); // getting attached intent data String address = h.getStringExtra("address"); // displaying selected product name LocationView.setText(address); */ Button ListViewScreen = (Button) findViewById(R.id.inputListView); //Listening to Button ListViewScreen.setOnClickListener(new View.OnClickListener() { public void onClick(View v) { //Switching to ListView Screen Intent i = new Intent(getApplicationContext(), ListViewActivity.class); startActivity(i); } } ); Button SelectedView = (Button) findViewById(R.id.inputListView); Intent i = getIntent(); // getting attached intent data String product = i.getStringExtra("product"); // displaying selected product name SelectedView.setText(product); Button ListView2Screen = (Button) findViewById(R.id.inputListView2); //Listening to Button ListView2Screen.setOnClickListener(new View.OnClickListener() { public void onClick(View v) { //Switching to ListView Screen Intent j = new Intent(getApplicationContext(), ListView2Activity.class); startActivity(j); } } ); Button SelectedView2 = (Button) findViewById(R.id.inputListView2); Intent j = getIntent(); // getting attached intent data String product2 = j.getStringExtra("product2"); // displaying selected product name SelectedView2.setText(product2); TextView Logout = (TextView) findViewById(R.id.linkLogout); // Listening to Log out Logout.setOnClickListener(new View.OnClickListener() { public void onClick(View arg0) { // Closing menu screen // Switching to Login Screen/closing register screen finish(); } }); } } Coding for listview class public class ListViewActivity extends ListActivity { @Override public void onCreate(Bundle savedInstanceState) { super.onCreate(savedInstanceState); // storing string resources into Array String[] traffic_condition = getResources().getStringArray(R.array.traffic_condition); // Binding resources Array to ListAdapter this.setListAdapter(new ArrayAdapter<String>(this, R.layout.listitem, R.id.listViewLayout, traffic_condition)); ListView lv = getListView(); // listening to single list item on click lv.setOnItemClickListener(new OnItemClickListener() { public void onItemClick(AdapterView<?> parent, View view, int position, long id) { // selected item String product = ((TextView) view).getText().toString(); // Launching new Activity on selecting single List Item Intent i = new Intent(getApplicationContext(), CheckinActivity.class); // sending data to new activity i.putExtra("product", product); startActivity(i); } }); } } Hope I made myself clear, I can provide a screen shot of my apps if it is required, thanks!

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  • seeking help with Chrome & Safari not rendering my table stretched to fit its contents...help?

    - by oompa_l
    I have an element on this web page I'm developing where I need my text to conform to the width of an image above it - whose width will always be different - think of captions. I have found numerous references to using a 1px table to force this width sizing behaviour. I am having problems, though with Safari and Chrome "seeing" this instruction - the text ends up as a marginally sized text box sitting behind the image. The problem, as I see it, has to do with the text and images sitting in div's nested within the table. I need the images to sit in a div because of some jquery script I'm using called cycle, which turns a group of images into a slideshow. The problem may have something to do with the script as well. In any case, I have tried a seeming infinite number of combination of floating left and clearing left on all all the divs, changing their positions and widths...nothing works. Anyone have any clues about how to broach this one? EDIT 1: ok, should I be editing my post or responding with answers? here's the url to see the problem I am having - http://friedmanstudios.ca/webdev/test8.html and the code: <div id="content" class="boxes"> <table> <tr> <td > <div id="imageFrame"> <a href="#" class="img" title="_MG_9786_fmt.jpeg"> <img src="images/_MG_9786_fmt.jpeg"/> </a> <a href="#" class="img" title="IMG_5169_fmt.jpeg"> <img src="indesign export/GFA-TEARSHEETS-100526-01-web-images/IMG_5169_fmt.jpeg"/> </a> <a href="#" class="img" title="IMG_5175_fmt.jpeg"> <img src="indesign export/GFA-TEARSHEETS-100526-01-web-images/IMG_5175_fmt.jpeg"/> </a> <a href="#" class="img" title="aerial_fmt.jpeg" width=""> <img src="indesign export/GFA-TEARSHEETS-100526-01-web-images/aerial_fmt.jpeg"/> </a> </div> <div id="cycleCtrl"> <div id="prev" class="pager"><a href="#">< Prev</a> </div> <div id="next" class="pager"><a href="#">Next ></a></div> <div id="pagerNav" class="pager"></div> </div> <div id="descController"> <img src="images/arrow.gif" name="arrow" width="5" height="10" id="arrow" /> <span id="projectName">Toronto Centre for the Arts </span> <br /> <div id="desc"> In the past eight years... </div> </div></td> <td width="90%"><!--push col 1 back--></td> </tr> </table> and the styles: #content { position: absolute; top: 250px; left: 275px; float: left; clear: both; } content table { float: left; width: 1px; } imageFrame { position: relative; float: left; clear: left; width: inherit; } desc { position: relative; clear: left; float: left; } descController { position:relative; padding-top:5px; padding-bottom:10px; clear: left; float: left; } descController div { height:0; overflow:hidden; -webkit-transition:all .5s ease; -moz-transition:all .5s ease; -o-transition:all .5s ease; transition:all .5s ease; padding-top:10px; margin-top: 10px; word-spacing: 0em; line-height: 16px; font-size: 12px; position: relative; float: left; clear: left; }

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  • tables wrapping to next line when width 100%

    - by jmo
    I'm encountering some weirdness with tables in css. The layout is fairly simple, a fixed-width nav bar on the left and the content on the right. When the content includes a table with a width of 100% the table ends up getting pushed down until it has room to take up the full width of the screen (instead of just the area to the right of the nav bar). If I remove the width=100% from the table's css, then it looks fine, but obviously the table doesn't grow to fill the space of the div. The problem is that i want the table to grow and shrink with the window but still stay in the bounds of its div. Thanks. Here's a simple example: <!DOCTYPE HTML PUBLIC "-//W3C//DTD HTML 4.01 Transitional//EN" "http://www.w3.org/TR/html4/loose.dtd"> <html> <head> <title>Test</title> <style type="text/css"> #content { padding-right:20px; background:white; overflow:hidden; margin:20px; } #content .column { position:relative; padding-bottom: 20010px; margin-bottom: -20000px; } #center { width:100%; padding-top:15px; } body { min-width:700px; } #left { width: 330px; padding: 0 10px; padding-top:10px; float:left; } .tableData { width:100%; } </style> </head> <body> <div id="content"> <div class="column" id="left"> <div> Some text goes in here<br/> some more text<br/> some more text<br/> some more text<br/> some more text<br/> some more text<br/> </div> </div> <div class="column" id="center"> Some text at the top; <hr/> <table class="tableData"> <thead> <tr><th>A</th><th>B</th><th>C</th></tr> </thead> <tbody> <tr> <td>A1 A1 A1 A1</td> <td>B1 B1 B1 B1</td> <td>C1 C1 C1 C1 C</td> </tr> <tr> <td>A2 A2 A2 A2 </td> <td>B2 B2 B2 B2 </td> <td>C2 C2 C2 C2</td> </tr> <tr> <td>A3 A3 A3 A3 A3 </td> <td>B3 B3 B3 B3 B3 </td> <td>C3 C3 C3 C3 C3</td> </tr> <tr> <td>A4 A4 A4 A4 A4</td> <td>B4 B4 B4 B4 B4</td> <td>C4 C4 C4 C4 C4</td> </tr> </tbody> </table> </div> </div> </body> </html>

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  • Where does ASP.NET Web API Fit?

    - by Rick Strahl
    With the pending release of ASP.NET MVC 4 and the new ASP.NET Web API, there has been a lot of discussion of where the new Web API technology fits in the ASP.NET Web stack. There are a lot of choices to build HTTP based applications available now on the stack - we've come a long way from when WebForms and Http Handlers/Modules where the only real options. Today we have WebForms, MVC, ASP.NET Web Pages, ASP.NET AJAX, WCF REST and now Web API as well as the core ASP.NET runtime to choose to build HTTP content with. Web API definitely squarely addresses the 'API' aspect - building consumable services - rather than HTML content, but even to that end there are a lot of choices you have today. So where does Web API fit, and when doesn't it? But before we get into that discussion, let's talk about what a Web API is and why we should care. What's a Web API? HTTP 'APIs' (Microsoft's new terminology for a service I guess)  are becoming increasingly more important with the rise of the many devices in use today. Most mobile devices like phones and tablets run Apps that are using data retrieved from the Web over HTTP. Desktop applications are also moving in this direction with more and more online content and synching moving into even traditional desktop applications. The pending Windows 8 release promises an app like platform for both the desktop and other devices, that also emphasizes consuming data from the Cloud. Likewise many Web browser hosted applications these days are relying on rich client functionality to create and manipulate the browser user interface, using AJAX rather than server generated HTML data to load up the user interface with data. These mobile or rich Web applications use their HTTP connection to return data rather than HTML markup in the form of JSON or XML typically. But an API can also serve other kinds of data, like images or other binary files, or even text data and HTML (although that's less common). A Web API is what feeds rich applications with data. ASP.NET Web API aims to service this particular segment of Web development by providing easy semantics to route and handle incoming requests and an easy to use platform to serve HTTP data in just about any content format you choose to create and serve from the server. But .NET already has various HTTP Platforms The .NET stack already includes a number of technologies that provide the ability to create HTTP service back ends, and it has done so since the very beginnings of the .NET platform. From raw HTTP Handlers and Modules in the core ASP.NET runtime, to high level platforms like ASP.NET MVC, Web Forms, ASP.NET AJAX and the WCF REST engine (which technically is not ASP.NET, but can integrate with it), you've always been able to handle just about any kind of HTTP request and response with ASP.NET. The beauty of the raw ASP.NET platform is that it provides you everything you need to build just about any type of HTTP application you can dream up from low level APIs/custom engines to high level HTML generation engine. ASP.NET as a core platform clearly has stood the test of time 10+ years later and all other frameworks like Web API are built on top of this ASP.NET core. However, although it's possible to create Web APIs / Services using any of the existing out of box .NET technologies, none of them have been a really nice fit for building arbitrary HTTP based APIs. Sure, you can use an HttpHandler to create just about anything, but you have to build a lot of plumbing to build something more complex like a comprehensive API that serves a variety of requests, handles multiple output formats and can easily pass data up to the server in a variety of ways. Likewise you can use ASP.NET MVC to handle routing and creating content in various formats fairly easily, but it doesn't provide a great way to automatically negotiate content types and serve various content formats directly (it's possible to do with some plumbing code of your own but not built in). Prior to Web API, Microsoft's main push for HTTP services has been WCF REST, which was always an awkward technology that had a severe personality conflict, not being clear on whether it wanted to be part of WCF or purely a separate technology. In the end it didn't do either WCF compatibility or WCF agnostic pure HTTP operation very well, which made for a very developer-unfriendly environment. Personally I didn't like any of the implementations at the time, so much so that I ended up building my own HTTP service engine (as part of the West Wind Web Toolkit), as have a few other third party tools that provided much better integration and ease of use. With the release of Web API for the first time I feel that I can finally use the tools in the box and not have to worry about creating and maintaining my own toolkit as Web API addresses just about all the features I implemented on my own and much more. ASP.NET Web API provides a better HTTP Experience ASP.NET Web API differentiates itself from the previous Microsoft in-box HTTP service solutions in that it was built from the ground up around the HTTP protocol and its messaging semantics. Unlike WCF REST or ASP.NET AJAX with ASMX, it’s a brand new platform rather than bolted on technology that is supposed to work in the context of an existing framework. The strength of the new ASP.NET Web API is that it combines the best features of the platforms that came before it, to provide a comprehensive and very usable HTTP platform. Because it's based on ASP.NET and borrows a lot of concepts from ASP.NET MVC, Web API should be immediately familiar and comfortable to most ASP.NET developers. Here are some of the features that Web API provides that I like: Strong Support for URL Routing to produce clean URLs using familiar MVC style routing semantics Content Negotiation based on Accept headers for request and response serialization Support for a host of supported output formats including JSON, XML, ATOM Strong default support for REST semantics but they are optional Easily extensible Formatter support to add new input/output types Deep support for more advanced HTTP features via HttpResponseMessage and HttpRequestMessage classes and strongly typed Enums to describe many HTTP operations Convention based design that drives you into doing the right thing for HTTP Services Very extensible, based on MVC like extensibility model of Formatters and Filters Self-hostable in non-Web applications  Testable using testing concepts similar to MVC Web API is meant to handle any kind of HTTP input and produce output and status codes using the full spectrum of HTTP functionality available in a straight forward and flexible manner. Looking at the list above you can see that a lot of functionality is very similar to ASP.NET MVC, so many ASP.NET developers should feel quite comfortable with the concepts of Web API. The Routing and core infrastructure of Web API are very similar to how MVC works providing many of the benefits of MVC, but with focus on HTTP access and manipulation in Controller methods rather than HTML generation in MVC. There’s much improved support for content negotiation based on HTTP Accept headers with the framework capable of detecting automatically what content the client is sending and requesting and serving the appropriate data format in return. This seems like such a little and obvious thing, but it's really important. Today's service backends often are used by multiple clients/applications and being able to choose the right data format for what fits best for the client is very important. While previous solutions were able to accomplish this using a variety of mixed features of WCF and ASP.NET, Web API combines all this functionality into a single robust server side HTTP framework that intrinsically understands the HTTP semantics and subtly drives you in the right direction for most operations. And when you need to customize or do something that is not built in, there are lots of hooks and overrides for most behaviors, and even many low level hook points that allow you to plug in custom functionality with relatively little effort. No Brainers for Web API There are a few scenarios that are a slam dunk for Web API. If your primary focus of an application or even a part of an application is some sort of API then Web API makes great sense. HTTP ServicesIf you're building a comprehensive HTTP API that is to be consumed over the Web, Web API is a perfect fit. You can isolate the logic in Web API and build your application as a service breaking out the logic into controllers as needed. Because the primary interface is the service there's no confusion of what should go where (MVC or API). Perfect fit. Primary AJAX BackendsIf you're building rich client Web applications that are relying heavily on AJAX callbacks to serve its data, Web API is also a slam dunk. Again because much if not most of the business logic will probably end up in your Web API service logic, there's no confusion over where logic should go and there's no duplication. In Single Page Applications (SPA), typically there's very little HTML based logic served other than bringing up a shell UI and then filling the data from the server with AJAX which means the business logic required for data retrieval and data acceptance and validation too lives in the Web API. Perfect fit. Generic HTTP EndpointsAnother good fit are generic HTTP endpoints that to serve data or handle 'utility' type functionality in typical Web applications. If you need to implement an image server, or an upload handler in the past I'd implement that as an HTTP handler. With Web API you now have a well defined place where you can implement these types of generic 'services' in a location that can easily add endpoints (via Controller methods) or separated out as more full featured APIs. Granted this could be done with MVC as well, but Web API seems a clearer and more well defined place to store generic application services. This is one thing I used to do a lot of in my own libraries and Web API addresses this nicely. Great fit. Mixed HTML and AJAX Applications: Not a clear Choice  For all the commonality that Web API and MVC share they are fundamentally different platforms that are independent of each other. A lot of people have asked when does it make sense to use MVC vs. Web API when you're dealing with typical Web application that creates HTML and also uses AJAX functionality for rich functionality. While it's easy to say that all 'service'/AJAX logic should go into a Web API and all HTML related generation into MVC, that can often result in a lot of code duplication. Also MVC supports JSON and XML result data fairly easily as well so there's some confusion where that 'trigger point' is of when you should switch to Web API vs. just implementing functionality as part of MVC controllers. Ultimately there's a tradeoff between isolation of functionality and duplication. A good rule of thumb I think works is that if a large chunk of the application's functionality serves data Web API is a good choice, but if you have a couple of small AJAX requests to serve data to a grid or autocomplete box it'd be overkill to separate out that logic into a separate Web API controller. Web API does add overhead to your application (it's yet another framework that sits on top of core ASP.NET) so it should be worth it .Keep in mind that MVC can generate HTML and JSON/XML and just about any other content easily and that functionality is not going away, so just because you Web API is there it doesn't mean you have to use it. Web API is not a full replacement for MVC obviously either since there's not the same level of support to feed HTML from Web API controllers (although you can host a RazorEngine easily enough if you really want to go that route) so if you're HTML is part of your API or application in general MVC is still a better choice either alone or in combination with Web API. I suspect (and hope) that in the future Web API's functionality will merge even closer with MVC so that you might even be able to mix functionality of both into single Controllers so that you don't have to make any trade offs, but at the moment that's not the case. Some Issues To think about Web API is similar to MVC but not the Same Although Web API looks a lot like MVC it's not the same and some common functionality of MVC behaves differently in Web API. For example, the way single POST variables are handled is different than MVC and doesn't lend itself particularly well to some AJAX scenarios with POST data. Code Duplication I already touched on this in the Mixed HTML and Web API section, but if you build an MVC application that also exposes a Web API it's quite likely that you end up duplicating a bunch of code and - potentially - infrastructure. You may have to create authentication logic both for an HTML application and for the Web API which might need something different altogether. More often than not though the same logic is used, and there's no easy way to share. If you implement an MVC ActionFilter and you want that same functionality in your Web API you'll end up creating the filter twice. AJAX Data or AJAX HTML On a recent post's comments, David made some really good points regarding the commonality of MVC and Web API's and its place. One comment that caught my eye was a little more generic, regarding data services vs. HTML services. David says: I see a lot of merit in the combination of Knockout.js, client side templates and view models, calling Web API for a responsive UI, but sometimes late at night that still leaves me wondering why I would no longer be using some of the nice tooling and features that have evolved in MVC ;-) You know what - I can totally relate to that. On the last Web based mobile app I worked on, we decided to serve HTML partials to the client via AJAX for many (but not all!) things, rather than sending down raw data to inject into the DOM on the client via templating or direct manipulation. While there are definitely more bytes on the wire, with this, the overhead ended up being actually fairly small if you keep the 'data' requests small and atomic. Performance was often made up by the lack of client side rendering of HTML. Server rendered HTML for AJAX templating gives so much better infrastructure support without having to screw around with 20 mismatched client libraries. Especially with MVC and partials it's pretty easy to break out your HTML logic into very small, atomic chunks, so it's actually easy to create small rendering islands that can be used via composition on the server, or via AJAX calls to small, tight partials that return HTML to the client. Although this is often frowned upon as to 'heavy', it worked really well in terms of developer effort as well as providing surprisingly good performance on devices. There's still plenty of jQuery and AJAX logic happening on the client but it's more manageable in small doses rather than trying to do the entire UI composition with JavaScript and/or 'not-quite-there-yet' template engines that are very difficult to debug. This is not an issue directly related to Web API of course, but something to think about especially for AJAX or SPA style applications. Summary Web API is a great new addition to the ASP.NET platform and it addresses a serious need for consolidation of a lot of half-baked HTTP service API technologies that came before it. Web API feels 'right', and hits the right combination of usability and flexibility at least for me and it's a good fit for true API scenarios. However, just because a new platform is available it doesn't meant that other tools or tech that came before it should be discarded or even upgraded to the new platform. There's nothing wrong with continuing to use MVC controller methods to handle API tasks if that's what your app is running now - there's very little to be gained by upgrading to Web API just because. But going forward Web API clearly is the way to go, when building HTTP data interfaces and it's good to see that Microsoft got this one right - it was sorely needed! Resources ASP.NET Web API AspConf Ask the Experts Session (first 5 minutes) © Rick Strahl, West Wind Technologies, 2005-2012Posted in Web Api   Tweet !function(d,s,id){var js,fjs=d.getElementsByTagName(s)[0];if(!d.getElementById(id)){js=d.createElement(s);js.id=id;js.src="//platform.twitter.com/widgets.js";fjs.parentNode.insertBefore(js,fjs);}}(document,"script","twitter-wjs"); (function() { var po = document.createElement('script'); po.type = 'text/javascript'; po.async = true; po.src = 'https://apis.google.com/js/plusone.js'; var s = document.getElementsByTagName('script')[0]; s.parentNode.insertBefore(po, s); })();

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  • Quick guide to Oracle IRM 11g: Server configuration

    - by Simon Thorpe
    Quick guide to Oracle IRM 11g index Welcome to the second article in this quick quide to Oracle IRM 11g. Hopefully you've just finished the first article which takes you through deploying the software onto a Linux server. This article walks you through the configuration of this new service and contains a subset of information from the official documentation and is focused on installing the server on Oracle Enterprise Linux. If you are planning to deploy on a non-Linux platform, you will need to reference the documentation for platform specific information. Contents Introduction Create IRM WebLogic Domain Starting the Admin Server and initial configuration Introduction In the previous article the database was prepared, the WebLogic Application Server installed and the files required for an IRM server installed. But we don't actually have a configured system yet. We need to now create a WebLogic Domain in which the IRM server will run, then configure some of the settings and crypography so that we can create a context and be ready to seal some content and test it all works. This article doesn't cover the configuration of SSL communication from client to server. This is quite a big topic and a separate article has been dedicated for this area. In these articles I also use the hostname, irm.company.internal to reference the IRM server and later on use the hostname irm.company.com in reference to the public facing service. Create IRM WebLogic Domain First step is creating the WebLogic domain, in a console switch to the newly created IRM installation folder as shown below and we will run the domain configuration wizard. [oracle@irm /]$ cd /oracle/middleware/Oracle_IRM/common/bin [oracle@irm bin]$ ./config.sh First thing the wizard will ask is if you wish to create a new or extend an existing domain. This guide is creating a standalone system so you should select to create a new domain. Next step is to choose what technologies from the Oracle ECM Suite you wish this domain to host. You are only interested in selecting the option "Oracle Information Rights Management". When you select this check box you will notice that it also selects "Oracle Enterprise Manager" and "Oracle JRF" as these are dependencies of the IRM server. You then need to specify where you wish to place the domain files. I usually just change the domain name from base_domain or irm_domain and leave the others with their defaults. Now the domain will have a single user initially and by default this user is called "weblogic". I usually change this account name to "sysadmin" or "administrator", but in this guide lets just accept the default. With respects to the next dialog, again for eval or dev reasons, leave the server startup mode as development. The JDK should also be automatically detected. We now need to provide details of the database. This guide is using the Oracle 11gR2 database and the settings I used can be seen in the image to the right. There is a lot of configuration that can now be done for the admin server, any managed servers and where the deployments reside. In this guide I am leaving all of these to their defaults so do not check any of the boxes. However I will on this blog be detailing later how you can go back and setup things such as automated startup of an IRM server which require changes to these default settings. But for now, lets leave it all alone and just click next. Now we are ready to install. Note that from this dialog you can scroll the left window and see there are going to be two servers created from the defaults. The AdminServer which is where you modify settings for the WebLogic Server and also hosts the Oracle Enterprise Manager for IRM which allows to monitor the IRM service performance and also make service related settings (which we shortly do below) and the IRM_server1 which hosts the actual IRM services themselves. So go right ahead and hit create, the process is pretty quick and usually under 10 minutes. When the domain creation ends, it will give you the URL to the admin server. It's worth noting this down and the URL is usually; http://irm.company.internal:7001 Starting the Admin Server and initial configuration First thing to do is to start the WebLogic Admin server and review the initial IRM server settings. In this guide we are going to run the Admin server and IRM server in console windows, in another article I will discuss running these as background services. So for now, start a console and run the Admin server by doing the following. cd /oracle/middleware/user_projects/domains/irm_domain/ ./startWebLogic.sh Wait for the server to start, you are looking for the following line to be reported in the console window. <BEA-00360><Server started in RUNNING mode> First step is configuring the IRM service via Enterprise Manager. Now that the Admin server is running you can point a browser at http://irm.company.internal:7001/em. Login with the username and password you supplied when you created the domain. In Enterprise Manager the IRM service administrator is able to make server wide configuration. However finding where to access the pages with these settings can be a bit of a challenge. After logging in on the left you'll see a tree containing elements of the Enterprise Manager farm Farm_irm_domain. Open up Content Management, then Information Rights Management and finally select the IRM node. On the right then select the IRM menu item, navigate to the Administration section and now we have four options, for now, we are just going to look at General Settings. The image on the right proves that a picture is worth a thousand words (or 113 in this case). The General Settings page allows you to set the cryptographic algorithms used for protecting sealed content. Unless you have a burning need to increase the key lengths or you need to comply to a regulation or government mandate, AES192 is a good start. You can change this later on without worry. The most important setting here we need to make is the Server URL. In this blog article I go over why this URL is so important, basically every single piece of content you protect with Oracle IRM is going to have this URL embedded in it, so if it's wrong or unresolvable, then nobody can open the secured documents. Note that in our environment we have yet to do any SSL configuration of the service. If you intend to build a server without SSL, then use http as the protocol instead of https. But I would recommend using SSL and setting this up is described in the next article. I would also probably up the device count from 1 to 3. This means that any user can retrieve rights to access content onto 3 computers at any one time. The default of 1 doesn't really make sense in development, evaluation nor even production environments and my experience is that 3 is a better number. Next step is to create the keystore for the IRM server. When a classification (called a context) is created, Oracle IRM generates a unique set of symmetric keys which are used to secure the content itself. These keys are then encrypted with a set of "wrapper" asymmetric cryptography keys which are stored externally to the server either in a Java Key Store or a HSM. These keys need to be generated and the following shows my commands and the resulting output. I have greyed out the responses from the commands so you can see the input a little easier. [oracle@irmsrv ~]$ cd /oracle/middleware/wlserver_10.3/server/bin/ [oracle@irmsrv bin]$ ./setWLSEnv.sh CLASSPATH=/oracle/middleware/patch_wls1033/profiles/default/sys_manifest_classpath/weblogic_patch.jar:/oracle/middleware/patch_ocp353/profiles/default/sys_manifest_classpath/weblogic_patch.jar:/usr/java/jdk1.6.0_18/lib/tools.jar:/oracle/middleware/wlserver_10.3/server/lib/weblogic_sp.jar:/oracle/middleware/wlserver_10.3/server/lib/weblogic.jar:/oracle/middleware/modules/features/weblogic.server.modules_10.3.3.0.jar:/oracle/middleware/wlserver_10.3/server/lib/webservices.jar:/oracle/middleware/modules/org.apache.ant_1.7.1/lib/ant-all.jar:/oracle/middleware/modules/net.sf.antcontrib_1.1.0.0_1-0b2/lib/ant-contrib.jar: PATH=/oracle/middleware/wlserver_10.3/server/bin:/oracle/middleware/modules/org.apache.ant_1.7.1/bin:/usr/java/jdk1.6.0_18/jre/bin:/usr/java/jdk1.6.0_18/bin:/usr/kerberos/bin:/usr/local/bin:/bin:/usr/bin:/home/oracle/bin Your environment has been set. [oracle@irmsrv bin]$ cd /oracle/middleware/user_projects/domains/irm_domain/config/fmwconfig/ [oracle@irmsrv fmwconfig]$ keytool -genkeypair -alias oracle.irm.wrap -keyalg RSA -keysize 2048 -keystore irm.jks Enter keystore password: Re-enter new password: What is your first and last name? [Unknown]: Simon Thorpe What is the name of your organizational unit? [Unknown]: Oracle What is the name of your organization? [Unknown]: Oracle What is the name of your City or Locality? [Unknown]: San Francisco What is the name of your State or Province? [Unknown]: CA What is the two-letter country code for this unit? [Unknown]: US Is CN=Simon Thorpe, OU=Oracle, O=Oracle, L=San Francisco, ST=CA, C=US correct? [no]: yes Enter key password for (RETURN if same as keystore password): At this point we now have an irm.jks in the directory /oracle/middleware/user_projects/domains/irm_domain/config/fmwconfig. The reason we store it here is this folder would be backed up as part of a domain backup. As with any cryptographic technology, DO NOT LOSE THESE KEYS OR THIS KEY STORE. Once you've sealed content against a context, the keys will be wrapped with these keys, lose these keys, and you can't get access to any secured content, pretty important. Now we've got the keys created, we need to go back to the IRM Enterprise Manager and set the location of the key store. Going back to the General Settings page in Enterprise Manager scroll down to Keystore Settings. Leave the type as JKS but change the location to; /oracle/Middleware/user_projects/domains/irm_domain/config/fmwconfig/irm.jks and hit Apply. The final step with regards to the key store is we need to tell the server what the password is for the Java Key Store so that it can be opened and the keys accessed. Once more fire up a console window and run these commands (again i've greyed out the clutter to see the commands easier). You will see dummy passed into the commands, this is because the command asks for a username, but in this instance we don't use one, hence the value dummy is passed and it isn't used. [oracle@irmsrv fmwconfig]$ cd /oracle/middleware/Oracle_IRM/common/bin/ [oracle@irmsrv bin]$ ./wlst.sh ... lots of settings fly by... Welcome to WebLogic Server Administration Scripting Shell Type help() for help on available commands wls:/offline>connect('weblogic','password','t3://irmsrv.us.oracle.com:7001') Connecting to t3://irmsrv.us.oracle.com:7001 with userid weblogic ... Successfully connected to Admin Server 'AdminServer' that belongs to domain 'irm_domain'. Warning: An insecure protocol was used to connect to the server. To ensure on-the-wire security, the SSL port or Admin port should be used instead. wls:/irm_domain/serverConfig>createCred("IRM","keystore:irm.jks","dummy","password") Location changed to domainRuntime tree. This is a read-only tree with DomainMBean as the root. For more help, use help(domainRuntime)wls:/irm_domain/serverConfig>createCred("IRM","key:irm.jks:oracle.irm.wrap","dummy","password") Already in Domain Runtime Tree wls:/irm_domain/serverConfig> At last we are now ready to fire up the IRM server itself. The domain creation created a managed server called IRM_server1 and we need to start this, use the following commands in a new console window. cd /oracle/middleware/user_projects/domains/irm_domain/bin/ ./startManagedWebLogic.sh IRM_server1 This will start up the server in the console, unlike the Admin server, you need to provide the username and password for the service to start. Enter in your weblogic username and password when prompted. You can change this behavior by putting the password into a boot.properties file, read more about this in the WebLogic Server documentation. Once running, wait until you see the line; <Notice><WebLogicServer><BEA-000360><Server started in RUNNING mode> At this point we can now login to the Oracle IRM Management Website at the URL. http://irm.company.internal:1600/irm_rights/ The server is just configured for HTTP at the moment, no SSL involved. Just want to ensure we can get a working system up and running. You should now see a login like the image on the right and you can now login using your weblogic username and password. The next article in this guide goes over adding SSL and now testing your server by actually adding a few users, sealing some content and opening this content as a user.

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  • Getting Started Building Windows 8 Store Apps with XAML/C#

    - by dwahlin
    Technology is fun isn’t it? As soon as you think you’ve figured out where things are heading a new technology comes onto the scene, changes things up, and offers new opportunities. One of the new technologies I’ve been spending quite a bit of time with lately is Windows 8 store applications. I posted my thoughts about Windows 8 during the BUILD conference in 2011 and still feel excited about the opportunity there. Time will tell how well it ends up being accepted by consumers but I’m hopeful that it’ll take off. I currently have two Windows 8 store application concepts I’m working on with one being built in XAML/C# and another in HTML/JavaScript. I really like that Microsoft supports both options since it caters to a variety of developers and makes it easy to get started regardless if you’re a desktop developer or Web developer. Here’s a quick look at how the technologies are organized in Windows 8: In this post I’ll focus on the basics of Windows 8 store XAML/C# apps by looking at features, files, and code provided by Visual Studio projects. To get started building these types of apps you’ll definitely need to have some knowledge of XAML and C#. Let’s get started by looking at the Windows 8 store project types available in Visual Studio 2012.   Windows 8 Store XAML/C# Project Types When you open Visual Studio 2012 you’ll see a new entry under C# named Windows Store. It includes 6 different project types as shown next.   The Blank App project provides initial starter code and a single page whereas the Grid App and Split App templates provide quite a bit more code as well as multiple pages for your application. The other projects available can be be used to create a class library project that runs in Windows 8 store apps, a WinRT component such as a custom control, and a unit test library project respectively. If you’re building an application that displays data in groups using the “tile” concept then the Grid App or Split App project templates are a good place to start. An example of the initial screens generated by each project is shown next: Grid App Split View App   When a user clicks a tile in a Grid App they can view details about the tile data. With a Split View app groups/categories are shown and when the user clicks on a group they can see a list of all the different items and then drill-down into them:   For the remainder of this post I’ll focus on functionality provided by the Blank App project since it provides a simple way to get started learning the fundamentals of building Windows 8 store apps.   Blank App Project Walkthrough The Blank App project is a great place to start since it’s simple and lets you focus on the basics. In this post I’ll focus on what it provides you out of the box and cover additional details in future posts. Once you have the basics down you can move to the other project types if you need the functionality they provide. The Blank App project template does exactly what it says – you get an empty project with a few starter files added to help get you going. This is a good option if you’ll be building an app that doesn’t fit into the grid layout view that you see a lot of Windows 8 store apps following (such as on the Windows 8 start screen). I ended up starting with the Blank App project template for the app I’m currently working on since I’m not displaying data/image tiles (something the Grid App project does well) or drilling down into lists of data (functionality that the Split App project provides). The Blank App project provides images for the tiles and splash screen (you’ll definitely want to change these), a StandardStyles.xaml resource dictionary that includes a lot of helpful styles such as buttons for the AppBar (a special type of menu in Windows 8 store apps), an App.xaml file, and the app’s main page which is named MainPage.xaml. It also adds a Package.appxmanifest that is used to define functionality that your app requires, app information used in the store, plus more. The App.xaml, App.xaml.cs and StandardStyles.xaml Files The App.xaml file handles loading a resource dictionary named StandardStyles.xaml which has several key styles used throughout the application: <Application x:Class="BlankApp.App" xmlns="http://schemas.microsoft.com/winfx/2006/xaml/presentation" xmlns:x="http://schemas.microsoft.com/winfx/2006/xaml" xmlns:local="using:BlankApp"> <Application.Resources> <ResourceDictionary> <ResourceDictionary.MergedDictionaries> <!-- Styles that define common aspects of the platform look and feel Required by Visual Studio project and item templates --> <ResourceDictionary Source="Common/StandardStyles.xaml"/> </ResourceDictionary.MergedDictionaries> </ResourceDictionary> </Application.Resources> </Application>   StandardStyles.xaml has style definitions for different text styles and AppBar buttons. If you scroll down toward the middle of the file you’ll see that many AppBar button styles are included such as one for an edit icon. Button styles like this can be used to quickly and easily add icons/buttons into your application without having to be an expert in design. <Style x:Key="EditAppBarButtonStyle" TargetType="ButtonBase" BasedOn="{StaticResource AppBarButtonStyle}"> <Setter Property="AutomationProperties.AutomationId" Value="EditAppBarButton"/> <Setter Property="AutomationProperties.Name" Value="Edit"/> <Setter Property="Content" Value="&#xE104;"/> </Style> Switching over to App.xaml.cs, it includes some code to help get you started. An OnLaunched() method is added to handle creating a Frame that child pages such as MainPage.xaml can be loaded into. The Frame has the same overall purpose as the one found in WPF and Silverlight applications - it’s used to navigate between pages in an application. /// <summary> /// Invoked when the application is launched normally by the end user. Other entry points /// will be used when the application is launched to open a specific file, to display /// search results, and so forth. /// </summary> /// <param name="args">Details about the launch request and process.</param> protected override void OnLaunched(LaunchActivatedEventArgs args) { Frame rootFrame = Window.Current.Content as Frame; // Do not repeat app initialization when the Window already has content, // just ensure that the window is active if (rootFrame == null) { // Create a Frame to act as the navigation context and navigate to the first page rootFrame = new Frame(); if (args.PreviousExecutionState == ApplicationExecutionState.Terminated) { //TODO: Load state from previously suspended application } // Place the frame in the current Window Window.Current.Content = rootFrame; } if (rootFrame.Content == null) { // When the navigation stack isn't restored navigate to the first page, // configuring the new page by passing required information as a navigation // parameter if (!rootFrame.Navigate(typeof(MainPage), args.Arguments)) { throw new Exception("Failed to create initial page"); } } // Ensure the current window is active Window.Current.Activate(); }   Notice that in addition to creating a Frame the code also checks to see if the app was previously terminated so that you can load any state/data that the user may need when the app is launched again. If you’re new to the lifecycle of Windows 8 store apps the following image shows how an app can be running, suspended, and terminated.   If the user switches from an app they’re running the app will be suspended in memory. The app may stay suspended or may be terminated depending on how much memory the OS thinks it needs so it’s important to save state in case the application is ultimately terminated and has to be started fresh. Although I won’t cover saving application state here, additional information can be found at http://msdn.microsoft.com/en-us/library/windows/apps/xaml/hh465099.aspx. Another method in App.xaml.cs named OnSuspending() is also included in App.xaml.cs that can be used to store state as the user switches to another application:   /// <summary> /// Invoked when application execution is being suspended. Application state is saved /// without knowing whether the application will be terminated or resumed with the contents /// of memory still intact. /// </summary> /// <param name="sender">The source of the suspend request.</param> /// <param name="e">Details about the suspend request.</param> private void OnSuspending(object sender, SuspendingEventArgs e) { var deferral = e.SuspendingOperation.GetDeferral(); //TODO: Save application state and stop any background activity deferral.Complete(); } The MainPage.xaml and MainPage.xaml.cs Files The Blank App project adds a file named MainPage.xaml that acts as the initial screen for the application. It doesn’t include anything aside from an empty <Grid> XAML element in it. The code-behind class named MainPage.xaml.cs includes a constructor as well as a method named OnNavigatedTo() that is called once the page is displayed in the frame.   /// <summary> /// An empty page that can be used on its own or navigated to within a Frame. /// </summary> public sealed partial class MainPage : Page { public MainPage() { this.InitializeComponent(); } /// <summary> /// Invoked when this page is about to be displayed in a Frame. /// </summary> /// <param name="e">Event data that describes how this page was reached. The Parameter /// property is typically used to configure the page.</param> protected override void OnNavigatedTo(NavigationEventArgs e) { } }   If you’re experienced with XAML you can switch to Design mode and start dragging and dropping XAML controls from the ToolBox in Visual Studio. If you prefer to type XAML you can do that as well in the XAML editor or while in split mode. Many of the controls available in WPF and Silverlight are included such as Canvas, Grid, StackPanel, and Border for layout. Standard input controls are also included such as TextBox, CheckBox, PasswordBox, RadioButton, ComboBox, ListBox, and more. MediaElement is available for rendering video or playing audio files. Some of the “common” XAML controls included out of the box are shown next:   Although XAML/C# Windows 8 store apps don’t include all of the functionality available in Silverlight 5, the core functionality required to build store apps is there with additional functionality available in open source projects such as Callisto (started by Microsoft’s Tim Heuer), Q42.WinRT, and others. Standard XAML data binding can be used to bind C# objects to controls, converters can be used to manipulate data during the data binding process, and custom styles and templates can be applied to controls to modify them. Although Visual Studio 2012 doesn’t support visually creating styles or templates, Expression Blend 5 handles that very well. To get started building the initial screen of a Windows 8 app you can start adding controls as mentioned earlier. Simply place them inside of the <Grid> element that’s included. You can arrange controls in a stacked manner using the StackPanel control, add a border around controls using the Border control, arrange controls in columns and rows using the Grid control, or absolutely position controls using the Canvas control. One of the controls that may be new to you is the AppBar. It can be used to add menu/toolbar functionality into a store app and keep the app clean and focused. You can place an AppBar at the top or bottom of the screen. A user on a touch device can swipe up to display the bottom AppBar or right-click when using a mouse. An example of defining an AppBar that contains an Edit button is shown next. The EditAppBarButtonStyle is available in the StandardStyles.xaml file mentioned earlier. <Page.BottomAppBar> <AppBar x:Name="ApplicationAppBar" Padding="10,0,10,0" AutomationProperties.Name="Bottom App Bar"> <Grid> <StackPanel x:Name="RightPanel" Orientation="Horizontal" Grid.Column="1" HorizontalAlignment="Right"> <Button x:Name="Edit" Style="{StaticResource EditAppBarButtonStyle}" Tag="Edit" /> </StackPanel> </Grid> </AppBar> </Page.BottomAppBar> Like standard XAML controls, the <Button> control in the AppBar can be wired to an event handler method in the MainPage.Xaml.cs file or even bound to a ViewModel object using “commanding” if your app follows the Model-View-ViewModel (MVVM) pattern (check out the MVVM Light package available through NuGet if you’re using MVVM with Windows 8 store apps). The AppBar can be used to navigate to different screens, show and hide controls, display dialogs, show settings screens, and more.   The Package.appxmanifest File The Package.appxmanifest file contains configuration details about your Windows 8 store app. By double-clicking it in Visual Studio you can define the splash screen image, small and wide logo images used for tiles on the start screen, orientation information, and more. You can also define what capabilities the app has such as if it uses the Internet, supports geolocation functionality, requires a microphone or webcam, etc. App declarations such as background processes, file picker functionality, and sharing can also be defined Finally, information about how the app is packaged for deployment to the store can also be defined. Summary If you already have some experience working with XAML technologies you’ll find that getting started building Windows 8 applications is pretty straightforward. Many of the controls available in Silverlight and WPF are available making it easy to get started without having to relearn a lot of new technologies. In the next post in this series I’ll discuss additional features that can be used in your Windows 8 store apps.

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  • How Expedia Made My New Bride Cry

    - by Lance Robinson
    Tweet this? Email Expedia and ask them to give me and my new wife our honeymoon? When Expedia followed up their failure with our honeymoon trip with a complete and total lack of acknowledgement of any responsibility for the problem and endless loops of explaining the issue over and over again - I swore that they would make it right. When they brought my new bride to tears, I got an immediate and endless supply of motivation. I hope you will help me make them make it right by posting our story on Twitter, Facebook, your blog, on Expedia itself, and when talking to your friends in person about their own travel plans.   If you are considering using them now for an important trip - reconsider. Short summary: We arrived early for a flight - but Expedia had made a mistake with the data they supplied to JetBlue and Emirates, which resulted in us not being able to check in (one leg of our trip was missing)!  At the time of this post, three people (myself, my wife, and an exceptionally patient JetBlue employee named Mary) each spent hours on the phone with Expedia.  I myself spent right at 3 hours (according to iPhone records), Lauren spent an hour and a half or so, and poor Mary was probably on the phone for a good 3.5 hours.  This is after 5 hours total at the airport.  If you add up our phone time, that is nearly 8 hours of phone time over a 5 hour period with little or no help, stall tactics (?), run-around, denial, shifting of blame, and holding. Details below (times are approximate): First, my wife and I were married yesterday - June 18th, the 3 year anniversary of our first date. She is awesome. She is the nicest person I have ever known, a ton of fun, absolutely beautiful in every way. Ok enough mushy - here are the dirty details. 2:30 AM - Early Check-in Attempt - we attempted to check-in for our flight online. Some sort of technology error on website, instructed to checkin at desk. 4:30 AM - Arrive at airport. Try to check-in at kiosk, get the same error. We got to the JetBlue desk at RDU International Airport, where Mary helped us. Mary discovered that the Expedia provided itinerary does not match the Expedia provided tickets. We are informed that when that happens American, JetBlue, and others that use the same software cannot check you in for the flight because. Why? Because the itinerary was missing a leg of our flight! Basically we were not shown in the system as definitely being able to make it home. Mary called Expedia and was put on hold by their automated system. 4:55 AM - Mary, myself, and my brand new bride all waited for about 25 minutes when finally I decided I would make a call myself on my iPhone while Mary was on the airport phone. In their automated system, I chose "make a new reservation", thinking they might answer a little more quickly than "customer service". Not surprisingly I was connected to an Expedia person within 1 minute. They informed me that they would have to forward me to a customer service specialist. I explained to them that we were already on hold for that and had been for nearly half an hour, that we were going on our honeymoon and that our flight would be leaving soon - could they please help us. "Yes, I will help you". I hand the phone to JetBlue Mary who explains the situation 3 or 4 times. Obviously I couldn't hear both ends of the conversation at this point, but the Expedia person explained what the problem was by stating exactly what Mary had just spent 15 minutes explaining. Mary calmly confirms that this is the problem, and asks Expedia to re-issue the itinerary. Expedia tells Mary that they'll have to transfer her to customer service. Mary asks for someone specific so that we get an answer this time, and goes on hold. Mary get's connected, explains the situation, and then Mary's connection gets terminated. 5:10 AM - Mary calls back to the Expedia automated system again, and we wait for about 5 minutes on hold this time before I pick up my iPhone and call Expedia again myself. Again I go to sales, a person picks up the phone in less than a minute. I explain the situation and let them know that we are now very close to missing our flight for our honeymoon, could they please help us. "Yes, I will help you". Again I give the phone to Mary who provides them with a call back number in case we get disconnected again and explains the situation again. More back and forth with Expedia doing nothing but repeating the same questions, Mary answering the questions with the same information she provided in the original explanation, and Expedia simply restating the problem. Mary again asks them to re-issue the itinerary, and explains that doing so will fix the problem. Expedia again repeats the problem instead of fixing it, and Mary's connection gets terminated. 5:20 AM - Mary again calls back to Expedia. My beautiful bride also calls on her own phone. At this point she is struggling to hold back her tears, stumbling through an explanation of all that has happened and that we are about to miss our flight. Please help us. "Yes, I will help". My beautiful bride's connection gets terminated. Ok, maybe this disconnection isn't an accident. We've now been disconnected 3 times on two different phones. 5:45 AM - I walk away and pleadingly beg a person to help me. They "escalate" the issue to "Rosy" (sp?) at Expedia. I go through the whole song and dance again with Rosy, who gives me the same treatment Mary was given. Rosy blames JetBlue for now having the correct data. Meanwhile Mary is on the phone with Emirates Air (the airline for the second leg of our trip), who agrees with JetBlue that Expedia's data isn't up to date. We are informed by two airport employees that issues like this with Expedia are not uncommon, and that the fix is simple. On the phone iwth Rosy, I ask her to re-issue the itinerary because we are about to miss our flight. She again explains the problem to me. At this point, I am standing at the window, pleading with Rosy to help us get to our honeymoon, watching our airplane. Then our airplane leaves without us. 6:03 AM - At this point we have missed our flight. Re-issuing the itinerary is no longer a solution. I ask Rosy to start from the beginning and work us up a new trip. She says that she cannot do that. She says that she needs to talk to JetBlue and Emirates and find out why we cannot check-in for our flight. I remind Rosy that our flight has already left - I just watched it taxi away - it no longer matters why (not to mention the fact that we already knew why, and have known why since 4:30 AM), and have known the solution since 4:30 AM. Rosy, can you please book a new trip? Yes, but it will cost $400. Excuse me? Now you can, but it will cost ME to fix your mistake? Rosy says that she can escalate the situation to her supervisor but that will take 1.5 hours. 6:15 AM - I told Rosy that if they had re-issued the itinerary as JetBlue asked (at 4:30 AM), my new wife and I might be on the airplane now instead of dealing with this on the phone and missing the beginning (and how much more?) of our honeymoon. Rosy said that it was not necessary to re-issue the itinerary. Out of curiosity, i asked Rosy if there was some financial burden on them to re-issue the itinerary. "No", said Rosy. I asked her if it was a large time burden on Expedia to re-issue the itinerary. "No", said Rosy. I directly asked Rosy: Why wouldn't Expedia have re-issued the itinerary when JetBlue asked? No answer. I asked Rosy: If you had re-issued the itinerary at 4:30, isn't it possible that I would be on that flight right now? She actually surprised me by answering "Yes" to that question. So I pointed out that it followed that Expedia was responsible for the fact that we missed out flight, and she immediately went into more about how the problem was with JetBlue - but now it was ALSO an Emirates Air problem as well. I tell Rosy to go ahead and escalate the issue again, and please call me back in that 1.5 hours (which how is about 1 hour and 10 minutes away). 6:30 AM - I start tweeting my frustration with iPhone. It's now pretty much impossible for us to make it to The Maldives by 3pm, which is the time at which we would need to arrive in order to be allowed service to the actual island where we are staying. Expedia has now given me the run-around for 2 hours, caused me to miss my flight, and worst of all caused my amazing new wife Lauren to miss our honeymoon. You think I was mad? No. Furious. Its ok to make mistakes - but to refuse to fix them and to ruin our honeymoon? No, not ok, Expedia. I swore right then that Expedia would make this right. 7:45 AM - JetBlue mary is still talking her tail off to other people in JetBlue and Emirates Air. Mary works it out so that if Expedia simply books a new trip, JetBlue and Emirates will both waive all the fees. Now we just have to convince Expedia to fix their mistake and get us on our way! Around this time Expedia Rosy calls me back! I inform her of the excellent work of JetBlue Mary - that JetBlue and Emirates both will waive the fees so Expedia can fix their mistake and get us going on our way. She says that she sees documentation of this in her system and that she needs to put me on hold "for 1 to 10 minutes" to talk to Emirates Air (why I'm not exactly sure). I say ok. 8:45 AM - After an hour on hold, Rosy comes on the line and asks me to hold more. I ask her to call me back. 9:35 AM - I put down the iPhone Twitter app and picks up the laptop. You think I made some noise with my iPhone? Heh 11:25 AM - Expedia follows me and sends a canned "We're sorry, DM us the details".  If you look at their Twitter feed, 16 out of the most recent 20 tweets are exactly the same canned response.  The other 4?  Ads.  Um - #MultiFAIL? To Expedia:  You now have had (as explained above) 8 hours of 3 different people explaining our situation, you know the email address of our Expedia account, you know my web blog, you know my Twitter address, you know my phone number.  You also know how upset you have made both me and my new bride by treating us with such a ... non caring, scripted, uncooperative, argumentative, and possibly even deceitful manner.  In the wise words of the great Kenan Thompson of SNL: "FIX IT!".  And no, I'm NOT going away until you make this right. Period. 11:45 AM - Expedia corporate office called.  The woman I spoke to was very nice and apologetic.  She listened to me tell the story again, she says she understands the problem and she is going to work to resolve it.  I don't have any details on what exactly that resolution might me, she said she will call me back in 20 minutes.  She found out about the problem via Twitter.  Thank you Twitter, and all of you who helped.  Hopefully social media will win my wife and I our honeymoon, and hopefully Expedia will encourage their customer service teams treat their customers properly. 12:22 PM - Spoke to Fran again from Expedia corporate office.  She has a flight for us tonight.  She is booking it now.  We will arrive at our honeymoon destination of beautiful Veligandu Island Resort only 1 day late.  She cannot confirm today, but she expects that Expedia will pay for the lost honeymoon night.  Thank you everyone for your help.  I will reflect more on this whole situation and confirm its resolution after our flight is 100% confirmed.  For now, I'm going to take a breather and go kiss my wonderful wife! 1:50 PM - Have not yet received the promised phone call.  We did receive an email with a new itinerary for a flight but the booking is not for specific seats, so there is no guarantee that my wife and I will be able to sit together.  With the original booking I carefully selected our seats for every segment of our trip.  I decided to call into the phone number that Fran from the Expedia corporate office gave me.  Its automated voice system identified itself as "Tier 3 Support".  I am currently still on hold with them, I have not gotten through to a human yet. 1:55 PM - Fran from Expedia called me back.  She confirmed us as booked.  She called the airlines to confirm.  Unfortunately, Expedia was unwilling or unable to allow us any type of seat selection.  It is possible that i won't get to sit next to the woman I married less than a day ago on our 40 total hours of flight time (there and back).  In addition, our seats could be the worst seats on the planes, with no reclining seat back or right next to the restroom.  Despite this fact (which in my opinion is huge), the horrible inconvenience, the hours at the airport, and the negative Internet publicity that Expedia is receiving, Expedia declined to offer us any kind of upgrade or to mark us as SFU (suitable for upgrade).  Since they didn't offer - I asked, and was rejected.  I am grateful to finally be heading in the right direction, but not only did Expedia horribly botch this job from the very beginning, they followed that botch job with near zero customer service, followed by a verbally apologetic but otherwise half-hearted resolution.  If this works out favorably for us, great.  If not - I'm not done making noise, Expedia.  You owe us, and I expect you to make it right.  You haven't quite done that yet. Thanks - Thank you to Twitter.  Thanks to all those who sympathize with us and helped us get the attention of Expedia, since three people (one of them an airline employee) using Expedia's normal channels of communication for many hours didn't help.  Thanks especially to my PowerShell and Sharepoint friends, my local friends, and those connectors who encouraged me and spread my story. 5:15 PM - Love Wins - After all this, Lauren and I are exhausted.  We both took a short nap, and when we woke up we talked about the last 24 hours.  It was a big, amazing, story-filled 24 hours.  I said that Expedia won, but Lauren said no.  She pointed out how lucky we are.  We are in love and married.  We have wonderful family and friends.  We are both hard-working successful people who love what they do.  We get to go to an amazing exotic destination for our honeymoon like Veligandu in The Maldives...  That's a lot of good.  Expedia didn't win.  This was (is) a big loss for Expedia.  It is a public blemish for all to see.  But Lauren and I did win, big time.  Expedia may not have made things right - but things are right for us.  Post in progress... I will relay any further comments (or lack of) from Expedia soon, as well as an update on confirmation of their repayment of our lost resort room rates.  I'll also post a picture of us on our honeymoon as soon as I can!

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  • Working with PivotTables in Excel

    - by Mark Virtue
    PivotTables are one of the most powerful features of Microsoft Excel.  They allow large amounts of data to be analyzed and summarized in just a few mouse clicks. In this article, we explore PivotTables, understand what they are, and learn how to create and customize them. Note:  This article is written using Excel 2010 (Beta).  The concept of a PivotTable has changed little over the years, but the method of creating one has changed in nearly every iteration of Excel.  If you are using a version of Excel that is not 2010, expect different screens from the ones you see in this article. A Little History In the early days of spreadsheet programs, Lotus 1-2-3 ruled the roost.  Its dominance was so complete that people thought it was a waste of time for Microsoft to bother developing their own spreadsheet software (Excel) to compete with Lotus.  Flash-forward to 2010, and Excel’s dominance of the spreadsheet market is greater than Lotus’s ever was, while the number of users still running Lotus 1-2-3 is approaching zero.  How did this happen?  What caused such a dramatic reversal of fortunes? Industry analysts put it down to two factors:  Firstly, Lotus decided that this fancy new GUI platform called “Windows” was a passing fad that would never take off.  They declined to create a Windows version of Lotus 1-2-3 (for a few years, anyway), predicting that their DOS version of the software was all anyone would ever need.  Microsoft, naturally, developed Excel exclusively for Windows.  Secondly, Microsoft developed a feature for Excel that Lotus didn’t provide in 1-2-3, namely PivotTables.  The PivotTables feature, exclusive to Excel, was deemed so staggeringly useful that people were willing to learn an entire new software package (Excel) rather than stick with a program (1-2-3) that didn’t have it.  This one feature, along with the misjudgment of the success of Windows, was the death-knell for Lotus 1-2-3, and the beginning of the success of Microsoft Excel. Understanding PivotTables So what is a PivotTable, exactly? Put simply, a PivotTable is a summary of some data, created to allow easy analysis of said data.  But unlike a manually created summary, Excel PivotTables are interactive.  Once you have created one, you can easily change it if it doesn’t offer the exact insights into your data that you were hoping for.  In a couple of clicks the summary can be “pivoted” – rotated in such a way that the column headings become row headings, and vice versa.  There’s a lot more that can be done, too.  Rather than try to describe all the features of PivotTables, we’ll simply demonstrate them… The data that you analyze using a PivotTable can’t be just any data – it has to be raw data, previously unprocessed (unsummarized) – typically a list of some sort.  An example of this might be the list of sales transactions in a company for the past six months. Examine the data shown below: Notice that this is not raw data.  In fact, it is already a summary of some sort.  In cell B3 we can see $30,000, which apparently is the total of James Cook’s sales for the month of January.  So where is the raw data?  How did we arrive at the figure of $30,000?  Where is the original list of sales transactions that this figure was generated from?  It’s clear that somewhere, someone must have gone to the trouble of collating all of the sales transactions for the past six months into the summary we see above.  How long do you suppose this took?  An hour?  Ten?  Probably. If we were to track down the original list of sales transactions, it might look something like this: You may be surprised to learn that, using the PivotTable feature of Excel, we can create a monthly sales summary similar to the one above in a few seconds, with only a few mouse clicks.  We can do this – and a lot more too! How to Create a PivotTable First, ensure that you have some raw data in a worksheet in Excel.  A list of financial transactions is typical, but it can be a list of just about anything:  Employee contact details, your CD collection, or fuel consumption figures for your company’s fleet of cars. So we start Excel… …and we load such a list… Once we have the list open in Excel, we’re ready to start creating the PivotTable. Click on any one single cell within the list: Then, from the Insert tab, click the PivotTable icon: The Create PivotTable box appears, asking you two questions:  What data should your new PivotTable be based on, and where should it be created?  Because we already clicked on a cell within the list (in the step above), the entire list surrounding that cell is already selected for us ($A$1:$G$88 on the Payments sheet, in this example).  Note that we could select a list in any other region of any other worksheet, or even some external data source, such as an Access database table, or even a MS-SQL Server database table.  We also need to select whether we want our new PivotTable to be created on a new worksheet, or on an existing one.  In this example we will select a new one: The new worksheet is created for us, and a blank PivotTable is created on that worksheet: Another box also appears:  The PivotTable Field List.  This field list will be shown whenever we click on any cell within the PivotTable (above): The list of fields in the top part of the box is actually the collection of column headings from the original raw data worksheet.  The four blank boxes in the lower part of the screen allow us to choose the way we would like our PivotTable to summarize the raw data.  So far, there is nothing in those boxes, so the PivotTable is blank.  All we need to do is drag fields down from the list above and drop them in the lower boxes.  A PivotTable is then automatically created to match our instructions.  If we get it wrong, we only need to drag the fields back to where they came from and/or drag new fields down to replace them. The Values box is arguably the most important of the four.  The field that is dragged into this box represents the data that needs to be summarized in some way (by summing, averaging, finding the maximum, minimum, etc).  It is almost always numerical data.  A perfect candidate for this box in our sample data is the “Amount” field/column.  Let’s drag that field into the Values box: Notice that (a) the “Amount” field in the list of fields is now ticked, and “Sum of Amount” has been added to the Values box, indicating that the amount column has been summed. If we examine the PivotTable itself, we indeed find the sum of all the “Amount” values from the raw data worksheet: We’ve created our first PivotTable!  Handy, but not particularly impressive.  It’s likely that we need a little more insight into our data than that. Referring to our sample data, we need to identify one or more column headings that we could conceivably use to split this total.  For example, we may decide that we would like to see a summary of our data where we have a row heading for each of the different salespersons in our company, and a total for each.  To achieve this, all we need to do is to drag the “Salesperson” field into the Row Labels box: Now, finally, things start to get interesting!  Our PivotTable starts to take shape….   With a couple of clicks we have created a table that would have taken a long time to do manually. So what else can we do?  Well, in one sense our PivotTable is complete.  We’ve created a useful summary of our source data.  The important stuff is already learned!  For the rest of the article, we will examine some ways that more complex PivotTables can be created, and ways that those PivotTables can be customized. First, we can create a two-dimensional table.  Let’s do that by using “Payment Method” as a column heading.  Simply drag the “Payment Method” heading to the Column Labels box: Which looks like this: Starting to get very cool! Let’s make it a three-dimensional table.  What could such a table possibly look like?  Well, let’s see… Drag the “Package” column/heading to the Report Filter box: Notice where it ends up…. This allows us to filter our report based on which “holiday package” was being purchased.  For example, we can see the breakdown of salesperson vs payment method for all packages, or, with a couple of clicks, change it to show the same breakdown for the “Sunseekers” package: And so, if you think about it the right way, our PivotTable is now three-dimensional.  Let’s keep customizing… If it turns out, say, that we only want to see cheque and credit card transactions (i.e. no cash transactions), then we can deselect the “Cash” item from the column headings.  Click the drop-down arrow next to Column Labels, and untick “Cash”: Let’s see what that looks like…As you can see, “Cash” is gone. Formatting This is obviously a very powerful system, but so far the results look very plain and boring.  For a start, the numbers that we’re summing do not look like dollar amounts – just plain old numbers.  Let’s rectify that. A temptation might be to do what we’re used to doing in such circumstances and simply select the whole table (or the whole worksheet) and use the standard number formatting buttons on the toolbar to complete the formatting.  The problem with that approach is that if you ever change the structure of the PivotTable in the future (which is 99% likely), then those number formats will be lost.  We need a way that will make them (semi-)permanent. First, we locate the “Sum of Amount” entry in the Values box, and click on it.  A menu appears.  We select Value Field Settings… from the menu: The Value Field Settings box appears. Click the Number Format button, and the standard Format Cells box appears: From the Category list, select (say) Accounting, and drop the number of decimal places to 0.  Click OK a few times to get back to the PivotTable… As you can see, the numbers have been correctly formatted as dollar amounts. While we’re on the subject of formatting, let’s format the entire PivotTable.  There are a few ways to do this.  Let’s use a simple one… Click the PivotTable Tools/Design tab: Then drop down the arrow in the bottom-right of the PivotTable Styles list to see a vast collection of built-in styles: Choose any one that appeals, and look at the result in your PivotTable:   Other Options We can work with dates as well.  Now usually, there are many, many dates in a transaction list such as the one we started with.  But Excel provides the option to group data items together by day, week, month, year, etc.  Let’s see how this is done. First, let’s remove the “Payment Method” column from the Column Labels box (simply drag it back up to the field list), and replace it with the “Date Booked” column: As you can see, this makes our PivotTable instantly useless, giving us one column for each date that a transaction occurred on – a very wide table! To fix this, right-click on any date and select Group… from the context-menu: The grouping box appears.  We select Months and click OK: Voila!  A much more useful table: (Incidentally, this table is virtually identical to the one shown at the beginning of this article – the original sales summary that was created manually.) Another cool thing to be aware of is that you can have more than one set of row headings (or column headings): …which looks like this…. You can do a similar thing with column headings (or even report filters). Keeping things simple again, let’s see how to plot averaged values, rather than summed values. First, click on “Sum of Amount”, and select Value Field Settings… from the context-menu that appears: In the Summarize value field by list in the Value Field Settings box, select Average: While we’re here, let’s change the Custom Name, from “Average of Amount” to something a little more concise.  Type in something like “Avg”: Click OK, and see what it looks like.  Notice that all the values change from summed totals to averages, and the table title (top-left cell) has changed to “Avg”: If we like, we can even have sums, averages and counts (counts = how many sales there were) all on the same PivotTable! Here are the steps to get something like that in place (starting from a blank PivotTable): Drag “Salesperson” into the Column Labels Drag “Amount” field down into the Values box three times For the first “Amount” field, change its custom name to “Total” and it’s number format to Accounting (0 decimal places) For the second “Amount” field, change its custom name to “Average”, its function to Average and it’s number format to Accounting (0 decimal places) For the third “Amount” field, change its name to “Count” and its function to Count Drag the automatically created field from Column Labels to Row Labels Here’s what we end up with: Total, average and count on the same PivotTable! Conclusion There are many, many more features and options for PivotTables created by Microsoft Excel – far too many to list in an article like this.  To fully cover the potential of PivotTables, a small book (or a large website) would be required.  Brave and/or geeky readers can explore PivotTables further quite easily:  Simply right-click on just about everything, and see what options become available to you.  There are also the two ribbon-tabs: PivotTable Tools/Options and Design.  It doesn’t matter if you make a mistake – it’s easy to delete the PivotTable and start again – a possibility old DOS users of Lotus 1-2-3 never had. We’ve included an Excel that should work with most versions of Excel, so you can download to practice your PivotTable skills. Download Our Practice Excel File Similar Articles Productive Geek Tips Magnify Selected Cells In Excel 2007Share Access Data with Excel in Office 2010Make Excel 2007 Print Gridlines In Workbook FileMake Excel 2007 Always Save in Excel 2003 FormatConvert Older Excel Documents to Excel 2007 Format TouchFreeze Alternative in AutoHotkey The Icy Undertow Desktop Windows Home Server – Backup to LAN The Clear & Clean Desktop Use This Bookmarklet to Easily Get Albums Use AutoHotkey to Assign a Hotkey to a Specific Window Latest Software Reviews Tinyhacker Random Tips Revo Uninstaller Pro Registry Mechanic 9 for Windows PC Tools Internet Security Suite 2010 PCmover Professional Ben & Jerry’s Free Cone Day, 3/23/10 New Stinger from McAfee Helps Remove ‘FakeAlert’ Threats Google Apps Marketplace: Tools & Services For Google Apps Users Get News Quick and Precise With Newser Scan for Viruses in Ubuntu using ClamAV Replace Your Windows Task Manager With System Explorer

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  • Setting up a new Silverlight 4 Project with WCF RIA Services

    - by Kevin Grossnicklaus
    Many of my clients are actively using Silverlight 4 and RIA Services to build powerful line of business applications.  Getting things set up correctly is critical to being to being able to take full advantage of the RIA services plumbing and when developers struggle with the setup they tend to shy away from the solution as a whole.  I’m a big proponent of RIA services and wanted to take the opportunity to share some of my experiences in setting up these types of projects.  In late 2010 I presented a RIA Services Master Class here in St. Louis, MO through my firm (ArchitectNow) and the information shared in this post was promised during that presentation. One other thing I want to mention before diving in is the existence of a number of other great posts on this subject.  I’ve learned a lot from many of them and wanted to call out a few of them.  The purpose of my post is to point out some of the gotchas that people get caught up on in the process but I would still encourage you to do as much additional research as you can to find the perfect setup for your needs. Here are a few additional blog posts and articles you should check out on the subject: http://msdn.microsoft.com/en-us/library/ee707351(VS.91).aspx http://adam-thompson.com/post/2010/07/03/Getting-Started-with-WCF-RIA-Services-for-Silverlight-4.aspx Technologies I don’t intend for this post to turn into a full WCF RIA Services tutorial but I did want to point out what technologies we will be using: Visual Studio.NET 2010 Silverlight 4.0 WCF RIA Services for Visual Studio 2010 Entity Framework 4.0 I also wanted to point out that the screenshots came from my personal development box which has a number of additional plug-ins and frameworks loaded so a few of the screenshots might not match 100% with what you see on your own machines. If you do not have Visual Studio 2010 you can download the express version from http://www.microsoft.com/express.  The Silverlight 4.0 tools and the WCF RIA Services components are installed via the Web Platform Installer (http://www.microsoft.com/web/download). Also, the examples given in this post are done in C#…sorry to you VB folks but the concepts are 100% identical. Setting up anew RIA Services Project This section will provide a step-by-step walkthrough of setting up a new RIA services project using a shared DLL for server side code and a simple Entity Framework model for data access.  All projects are created with the consistent ArchitectNow.RIAServices filename prefix and default namespace.  This would be modified to match your companies standards. First, open Visual Studio and open the new project window via File->New->Project.  In the New Project window, select the Silverlight folder in the Installed Templates section on the left and select “Silverlight Application” as your project type.  Verify your solution name and location are set appropriately.  Note that the project name we specified in the example below ends with .Client.  This indicates the name which will be given to our Silverlight project. I consider Silverlight a client-side technology and thus use this name to reflect that.  Click Ok to continue. During the creation on a new Silverlight 4 project you will be prompted with the following dialog to create a new web ASP.NET web project to host your Silverlight content.  As we are demonstrating the setup of a WCF RIA Services infrastructure, make sure the “Enable WCF RIA Services” option is checked and click OK.  Obviously, there are some other options here which have an effect on your solution and you are welcome to look around.  For our example we are going to leave the ASP.NET Web Application Project selected.  If you are interested in having your Silverlight project hosted in an MVC 2 application or a Web Site project these options are available as well.  Also, whichever web project type you select, the name can be modified here as well.  Note that it defaults to the same name as your Silverlight project with the addition of a .Web suffix. At this point, your full Silverlight 4 project and host ASP.NET Web Application should be created and will now display in your Visual Studio solution explorer as part of a single Visual Studio solution as follows: Now we want to add our WCF RIA Services projects to this same solution.  To do so, right-click on the Solution node in the solution explorer and select Add->New Project.  In the New Project dialog again select the Silverlight folder under the Visual C# node on the left and, in the main area of the screen, select the WCF RIA Services Class Library project template as shown below.  Make sure your project name is set appropriately as well.  For the sample below, we will name the project “ArchitectNow.RIAServices.Server.Entities”.   The .Server.Entities suffix we use is meant to simply indicate that this particular project will contain our WCF RIA Services entity classes (as you will see below).  Click OK to continue. Once you have created the WCF RIA Services Class Library specified above, Visual Studio will automatically add TWO projects to your solution.  The first will be an project called .Server.Entities (using our naming conventions) and the other will have the same name with a .Web extension.  The full solution (with all 4 projects) is shown in the image below.  The .Entities project will essentially remain empty and is actually a Silverlight 4 class library that will contain generated RIA Services domain objects.  It will be referenced by our front-end Silverlight project and thus allow for simplified sharing of code between the client and the server.   The .Entities.Web project is a .NET 4.0 class library into which we will put our data access code (via Entity Framework).  This is our server side code and business logic and the RIA Services plumbing will maintain a link between this project and the front end.  Specific entities such as our domain objects and other code we set to be shared will be copied automatically into the .Entities project to be used in both the front end and the back end. At this point, we want to do a little cleanup of the projects in our solution and we will do so by deleting the “Class1.cs” class from both the .Entities project and the .Entities.Web project.  (Has anyone ever intentionally named a class “Class1”?) Next, we need to configure a few references to make RIA Services work.  THIS IS A KEY STEP THAT CAUSES MANY HEADACHES FOR DEVELOPERS NEW TO THIS INFRASTRUCTURE! Using the Add References dialog in Visual Studio, add a project reference from the *.Client project (our Silverlight 4 client) to the *.Entities project (our RIA Services class library).  Next, again using the Add References dialog in Visual Studio, add a project reference from the *.Client.Web project (our ASP.NET host project) to the *.Entities.Web project (our back-end data services DLL).  To get to the Add References dialog, simply right-click on the project you with to add a reference to in the Visual Studio solution explorer and select “Add Reference” from the resulting context menu.  You will want to make sure these references are added as “Project” references to simplify your future debugging.  To reiterate the reference direction using the project names we have utilized in this example thus far:  .Client references .Entities and .Client.Web reference .Entities.Web.  If you have opted for a different naming convention, then the Silverlight project must reference the RIA Services Silverlight class library and the ASP.NET host project must reference the server-side class library. Next, we are going to add a new Entity Framework data model to our data services project (.Entities.Web).  We will do this by right clicking on this project (ArchitectNow.Server.Entities.Web in the above diagram) and selecting Add->New Project.  In the New Project dialog we will select ADO.NET Entity Data Model as in the following diagram.  For now we will call this simply SampleDataModel.edmx and click OK. It is worth pointing out that WCF RIA Services is in no way tied to the Entity Framework as a means of accessing data and any data access technology is supported (as long as the server side implementation maps to the RIA Services pattern which is a topic beyond the scope of this post).  We are using EF to quickly demonstrate the RIA Services concepts and setup infrastructure, as such, I am not providing a database schema with this post but am instead connecting to a small sample database on my local machine.  The following diagram shows a simple EF Data Model with two tables that I reverse engineered from a local data store.   If you are putting together your own solution, feel free to reverse engineer a few tables from any local database to which you have access. At this point, once you have an EF data model generated as an EDMX into your .Entites.Web project YOU MUST BUILD YOUR SOLUTION.  I know it seems strange to call that out but it important that the solution be built at this point for the next step to be successful.  Obviously, if you have any build errors, these must be addressed at this point. At this point we will add a RIA Services Domain Service to our .Entities.Web project (our server side code).  We will need to right-click on the .Entities.Web project and select Add->New Item.  In the Add New Item dialog, select Domain Service Class and verify the name of your new Domain Service is correct (ours is called SampleService.cs in the image below).  Next, click "Add”. After clicking “Add” to include the Domain Service Class in the selected project, you will be presented with the following dialog.  In it, you can choose which entities from the selected EDMX to include in your services and if they should be allowed to be edited (i.e. inserted, updated, or deleted) via this service.  If the “Available DataContext/ObjectContext classes” dropdown is empty, this indicates you have not yes successfully built your project after adding your EDMX.  I would also recommend verifying that the “Generate associated classes for metadata” option is selected.  Once you have selected the appropriate options, click “OK”. Once you have added the domain service class to the .Entities.Web project, the resulting solution should look similar to the following: Note that in the solution you now have a SampleDataModel.edmx which represents your EF data mapping to your database and a SampleService.cs which will contain a large amount of generated RIA Services code which RIA Services utilizes to access this data from the Silverlight front-end.  You will put all your server side data access code and logic into the SampleService.cs class.  The SampleService.metadata.cs class is for decorating the generated domain objects with attributes from the System.ComponentModel.DataAnnotations namespace for validation purposes. FINAL AND KEY CONFIGURATION STEP!  One key step that causes significant headache to developers configuring RIA Services for the first time is the fact that, when we added the EDMX to the .Entities.Web project for our EF data access, a connection string was generated and placed within a newly generated App.Context file within that project.  While we didn’t point it out at the time you can see it in the image above.  This connection string will be required for the EF data model to successfully locate it’s data.  Also, when we added the Domain Service class to the .Entities.Web project, a number of RIA Services configuration options were added to the same App.Config file.   Unfortunately, when we ultimately begin to utilize the RIA Services infrastructure, our Silverlight UI will be making RIA services calls through the ASP.NET host project (i.e. .Client.Web).  This host project has a reference to the .Entities.Web project which actually contains the code so all will pass through correctly EXCEPT the fact that the host project will utilize it’s own Web.Config for any configuration settings.  For this reason we must now merge all the sections of the App.Config file in the .Entities.Web project into the Web.Config file in the .Client.Web project.  I know this is a bit tedious and I wish there were a simpler solution but it is required for our RIA Services Domain Service to be made available to the front end Silverlight project.  Much of this manual merge can be achieved by simply cutting and pasting from App.Config into Web.Config.  Unfortunately, the <system.webServer> section will exist in both and the contents of this section will need to be manually merged.  Fortunately, this is a step that needs to be taken only once per solution.  As you add additional data structures and Domain Services methods to the server no additional changes will be necessary to the Web.Config. Next Steps At this point, we have walked through the basic setup of a simple RIA services solution.  Unfortunately, there is still a lot to know about RIA services and we have not even begun to take advantage of the plumbing which we just configured (meaning we haven’t even made a single RIA services call).  I plan on posting a few more introductory posts over the next few weeks to take us to this step.  If you have any questions on the content in this post feel free to reach out to me via this Blog and I’ll gladly point you in (hopefully) the right direction. Resources Prior to closing out this post, I wanted to share a number or resources to help you get started with RIA services.  While I plan on posting more on the subject, I didn’t invent any of this stuff and wanted to give credit to the following areas for helping me put a lot of these pieces into place.   The books and online resources below will go a long way to making you extremely productive with RIA services in the shortest time possible.  The only thing required of you is the dedication to take advantage of the resources available. Books Pro Business Applications with Silverlight 4 http://www.amazon.com/Pro-Business-Applications-Silverlight-4/dp/1430272074/ref=sr_1_2?ie=UTF8&qid=1291048751&sr=8-2 Silverlight 4 in Action http://www.amazon.com/Silverlight-4-Action-Pete-Brown/dp/1935182374/ref=sr_1_1?ie=UTF8&qid=1291048751&sr=8-1 Pro Silverlight for the Enterprise (Books for Professionals by Professionals) http://www.amazon.com/Pro-Silverlight-Enterprise-Books-Professionals/dp/1430218673/ref=sr_1_3?ie=UTF8&qid=1291048751&sr=8-3 Web Content RIA Services http://channel9.msdn.com/Blogs/RobBagby/NET-RIA-Services-in-5-Minutes http://silverlight.net/riaservices/ http://www.silverlight.net/learn/videos/all/net-ria-services-intro/ http://www.silverlight.net/learn/videos/all/ria-services-support-visual-studio-2010/ http://channel9.msdn.com/learn/courses/Silverlight4/SL4BusinessModule2/SL4LOB_02_01_RIAServices http://www.myvbprof.com/MainSite/index.aspx#/zSL4_RIA_01 http://channel9.msdn.com/blogs/egibson/silverlight-firestarter-ria-services http://msdn.microsoft.com/en-us/library/ee707336%28v=VS.91%29.aspx Silverlight www.silverlight.net http://msdn.microsoft.com/en-us/silverlight4trainingcourse.aspx http://channel9.msdn.com/shows/silverlighttv

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